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A TEXT BOOK 
OF 
FIRE ASSAYING 


EDWARD E. BUGBEE 








© Raymond Pettibon 


RESEARCH LIBRARY 
THE GETTY RESEARCH INSTITUTE 


JOHN MOORE ANDREAS COLOR CHEMISTRY LIBRARY FOUNDATION 












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ex LEX T- BOOK 


OF 


RE ASSAYING 


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A TEXT-BOOK 


OF 


FIRE ASSAYING 


EDWARD E. BUGBEE 


Assistant Professor of Mining Engineering and Metallurgy, 
Massachusetts Institute of Technology. 


Printed for use in Gofinection with 


THE COURSE IN FIRE ASSAYING AT THE 


MASSACHUSETTS INSTITUTE OF TECHNOLOGY. 


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Copyright 1915, rele 
. By EDWARD E. BUGBEE. 


PRESS OF 





BOSTON, MASS. Ne 





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THE GETTY RESEARCH 
INS UW |BRARY Sem 


CONTENTS 


CHAPTER I 
AssAY REAGENTS AND FUSION PRODUCTS .......2.+eee Silos 
. Definitions. Reagents. Chemical Reactions of Reagents. Fusion 
Products. 
é 
| CHAPTER II 
. FURNACES AND MrrauuurcicaL Cuay Goops ........4.4... 


Crucible Furnaces. Muffle Furnaces. Fuels. Coal Furnace. Coke 
Furnace. Gasoline Furnace. Gas Furnace. Crude Oil Furnace. 
Furnace Repairs. Muffles. Crucibles. Scorifiers. Roasting Dishes, 
etc. Testing Crucibles. 


CHAPTER III 






en eg ig oe pence as amo e wow he Sh wt ewe 


Definitions. Moisture Sample. Sampling Operations. Theory of 
Ore Sampling. Duplicate Sampling. Sampling Ore Containing 
Malleable Minerals. 


SHAPLER.LY 


UMIIMMERRCTOMVVCRSIPER ¢ 4 . 6 4k ke we ee ke we 8s 


Flux Balance. Pulp Balance. Button Balance. Theory of the 
Balance. Directions for use of Balance. Weighing by Equal Swings. 
Weighing by No Deflection. Weighing by Substitution. Check 
Weighing. Adjusting and Testing an Assay Balance. Weights. 
Calibration of Weights. Riders. Testing Riders. 


CHAPTER V. 


Cupellation. Assay of Lead Bullion. Loss of Silver in Cupelling. 
Effect of Temperature. Protective Action of Silver on Gold. Influence 
of Impurities. Indications of Metals Present. Testing Cupels. Re- 
tention of Base Metals. Portland Cement and Magnesia Cupels. 


9-20 


21-36 


37-50 


51-67 


iv 
CHAPTER VI 


PARTING °°) 2 soe See ee ee eee 


General Statement. Parting in Porcelain Capsules. Inquartation. 
Parting in Flasks. 


CHAPTER VII 


Tue ScoriFICATION ASSAY 


@ @ @) -@. “re. Sree, 8 Ry i ie ee 


General Statement. Solubility of Metallic Oxides in Litharge. Heat 
of Formation of Metallic Oxides. Ignition Temperature of Metallic 
Sulphides. Assay Procedure for Ores. Chemical Reactions. Color 
of Scorifiers. Assaying Granulated Lead. Scorification Assay for 
Gold. Scorification Assay of Copper Matte. Table of Scorification 
Charges for Different Ores. 


CHAPTER VIII 


Tar Cructpir Assay 


Theory of the Crucible Assay. Classification of Ores. Crucible 
Slags.‘ Classification of Silicates. Fluidity of Slags. Acid and Basic 
Slags. Mixed Silicates. Notes on the Fusibility of Silicates. Slags 
for Class 1 Siliceous Ores. Slags for Class 1 Basic Ores. Reducing and 
Oxidizing. Reducing Power of Minerals. Slags for Class 2 Ores. 
Action of Borax in Slags. The Cover. Testing Reagents. Assay of 
Class 1 Ores. Methods for Assay of Class 2 Ores. The Niter Method. 
Preliminary Fusion. Estimating Reducing Power. Regular Fusion. 
The Iron Method. The Roasting Method. Assay of Class 3 Ores. 
Determining the Oxidizing Power of Ores. 


CHAPTER IX 


Sprciat MetHops of AssAy .....2.... | ee 


The Assay of Telluride Ores. Crucible Methods for the Assay of 
Ores and Products High in Copper. The Assay of Antimonial Gold 
Ores. The Assay of Auriferouis Tinstone. The Corrected Assay. 


CHAPTER X 


THe Assay OF BULLION o o e . cde a Pee 4 . e . é * . . * . . 


Definition. * Weights. Sampling Bullion. The Assay of Lead 
Bullion. The Assay of Copper Bullion. Atl Fire Method. Crucible 
Method. Mercury—Stulphuric Acid Combination Method. Nitric 
Acid Combination Method. The Assay of Doré Bullion. U.S. Mint 
Assay of Gold Bullion. | 


68-72 


73-82 


83-111 


. 112-118 


119=132 


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CHAPTER XI 






OC ROL UTIONS we es ek wee 


_ Evaporation in Lead Tray. Evaporation with Litharge. Precipita- 
tion by Zinc and Lead Acetate. Precipitation as Sulphide. Precipi- 
tation by Cement Copper. Precipitation by Silver Nitrate. Precipi- 
tation by a Copper Salt. Electrolytic Precipitation. Colorimetric 
~ Method. 


CHAPTER XII 


a BO) rc 


_ General Statement. Lead Ores. Accuracy and Limitations of 
Method. Quantity of Ore and Reagents Used. Manipulation of 
Assay. Influence of Other Metals. Procedure for Assay. Assay of 
~Slags. ; 


e e e ° ° ° ° e e e ° e ° e e ° ° e e e e e e e . 





133-139 


140-145 


147-150 








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A TEXT-BOOK OF FIRE ASSAYING. 


GHAPTER. I. 


ASSAY REAGENTS AND FUSION PRODUCTS. 


Assaying is a branch of analytical chemistry, generally defined as 
the quantitative estimation of the metals in ores and furnace prod- 
ucts. Inthe Western part of the United States, the term is employed 
to include the determination of all the constituents, both metallic 
and non-metallic, of ores and metallurgical products. 


Fire Assaying is the quantitative determination of metals in ores 
and metallurgical products by means of heat and dry reagents. This 
involves the separation of the metal from the other constituents of 
the ore and its weighing in a state of purity. 

The reagents used in fire assaying may be classified as fluxes; acid, 
basic or neutral, and as oxidizing, reducing, sulphurizing or desulphur- 
izing agents. Some reagents have simply one property, as for in- 
stance silica, an acid flux, others have several different properties, as 
litharge, a basic flux but also an oxidizing and desulphurizing agent. 

A flux is something, which, if added to a body infusible by itself or 
fusible only with difficulty, will cause it to fuse at a lower temperature 
than it would have done alone. For instance, quartz by itself is’ 
fusible only at a very high temperature, but by adding some sodium 
carbonate, to the pulverized quartz it can be fused at a temperature 
easily obtained in the assay furnace. 

The student should remember that to aid in the fusion of an acid 
substance, a basic flux such as litharge, sodium carbonate, limestone, 
or iron oxide should be added, for a basic substance an acid flux such 
as silica or borax should be used. 


The principal fluxes used in assaying follow:— 


Silica, SiO», is an acid flux and the strongest one we have. It 
combines with the metal oxides to form silicates which are the founda- 
tion of almost all of our slags. It is used as a flux when the ore is 
deficient in silica and serves to protect the crucibles and scorifiers 
from the corrosive action of litharge. Care must be taken to avoid 


2 


an excess of silica, as too much of it will cause trouble and losses of 
precious metals by slagging or by the formation of a matte. Silica 
melts at about 1625° C. to an extremely viscous liquid. (Day «& 
Shepherd). It should be obtained in the pulverized form. 


Glass is used by some in place of silica. Ordinary window glass, 
a silicate of lime and the alkalies with the silica in excess, is best. Its 
acid excess is always doubtful and so is not commonly used. If used, 
a blank assay should be run on each new lot to insure against intro- 
ducing precious metals into the assay in this way. Its chief advan- 
tage is that 5 or 10 grams too much glass will ordinarily do no harm 
in a fusion whereas 5 or 10 grams of silica in excess might spoil it. 


Borax, Na:B,0;,10H:O, acts as an acid flux. It contains a large 
amount (47 per cent) of water which is given off on heating. During 
the heating, the borax swells to more than twice its original bulk, 
and if an excess of it is used, and especially if not thoroughly mixed 
with the charge, it may force part of the charge out of the crucible. 
It should never be used for scorification work. 


Borax Glass, Na2,B,O;, is an active, readily fusible, acid flux. It 
is made by fusing borax in a crucible and pouring the fused mass on 


a clean iron or brick surface. It should be crushed to pass a 16 or 


20 mesh screen before using. Crystaline borax glass melts at 742° C. 
Finely divided amorphous borax glass begins to sinter at 490°-500° C. 
It is extremely viscous when melted. 

Its rational formula Na,O, 2B.03 indicates an excess of acid and 
experiment proves this to be the case. It is an excellent flux for all 
the metallic oxides, and fusing as it does at a low temperature, it 
helps to facilitate the slagging of the ore. Either borax or borax 
glass is used in almost-every flux mixture. It should be used in prefer- 
ence to silica as a flux for ime, magnesia, iron, manganese and zinc 
oxides. 

- Borax glass cannot always be used in preference to borax, as it is 
more violent in its action and causes much boiling of the charge. 
The fine material especially takes on moisture from the air and tends 


to cake. It is sometimes used as a cover on top of a crucible charge, — 


especially in muffle fusions. Here it melts before the rest of the charge 
and prevents loss of the fine ore by “dusting.” 


Sodium bicarbonate, NaHCOs, is the most common of the alkali 


carbonate fluxes. It is decomposed on heating to 276° C. forming the 


normal carbonate as shown by the following reaction :— 


2NaHCO; = NaeCO; + H,O + COz 


Pre 


3 





It acts as a basic flux, a desulphurizing agent and in some cases 
as an oxidizing agent. It is the cheapest of the alkali carbonates, 
and, as it is readily obtained pure, and does not deliquesce, it is pre- 
ferred by many to the normal carbonate. 


Sodium carbonate, Na.COs, (normal carbonate) melts at 852° C. 
: It occurs crystallized, (sal-soda) containing more than half its weight 
of water, and in this form, is not at all suited for a flux. In the 
' anhydrous form, it is somewhat stronger, weight for weight, than the 
bicarbonate, but unfortunately it tends to absorb water from the air 
| and is therefore unsatisfactory for use in some climates. The com- 
mercial normal carbonate of this country is made by the Solvay pro- 
cess from the bicarbonate and should be very pure. The variety 
known by the trade as, 58% dense, soda-ash has been found entirely 
satisfactory for assay purposes and is but little affected by atmospheric 
moisture. 

It acts as a basic flux, a desulphurizing agent, and in some cases, an 
oxidizing agent. Heated in the presence of silica, it breaks up giving 
off CO, and the Na,O combines with the silica forming silicates, as 
for example :— | 


NasCO; a SiOs = NasSi0; a CO. 




















| Many of these silicates are readily fusible and for this reason soda 
is used in nearly every fusion. 


Potassium carbonate, K,COs, fused at 894° C. and is also a basie 
flux. Its action is in all ways similar to sodium carbonate. 

A mixture of sodium and potassium carbonates fuses at a lower 
temperature than either one alone, due to the formation of a double 
salt and therefore the mixture is used whenever it is desired to main- 
tain a low temperature during the assay. The lead assay is a case in 
point. 7 

Potassium carbonate should be kept in an air tight receptacle as 
otherwise it takes on water from the air and forms a hard cake. 


Litharge, PbO, (92.83% lead) is a readily fusible basic flux. It 
~ acts also as an oxidizing and desulphurizing agent and on being re- 
duced it supplies the lead necessary for the collection of the gold and 
silver. It melts at 883°C. (Mostowitsch). 
- Litharge begins to combine with silica at 600° C. forming lead 
silicates which are pasty at this temperature. 


2PbO + $10. = Pb.SiO, (monosilicate). This silicate is fusible 


Fa ipl 


— 


+ 


at a low temperature 746° C. and is as fluid as water. If the pro- 
portion of silica be raised above that of the tri-silicate, the mixture 
becomes less easily fusible and is decidedly viscous when fused. 
Litharge has such a strong affinity for silica that if enough is not sup- 
plied to it in the crucible charge, it will attack the acid material of the 
crucible itself and if left long enough will eat a hole through it. 

Litharge readily gives up its oxygen if heated with carbon, carbonic 
oxide, hydrogen, sulphur, metallic sulphides, iron, ete. The reaction 
with carbon begins at about 550°C. It thus acts as an oxidizing, 
and in the presence of sulphur, as a desulphurizing agent :— 

2PbO + C = CO, + 2Pb (oxidizing) 
3PbO + ZnS = ZnO + SO. + 3Pb  (desulphurizing and oxidiz- 
ing) . 
The liberated lead is then available for the collection of the gold and 
silver. 

Lead silicates do not readily give up their lead to carbonaceous and 
sulphurous reducing agents. The higher the proportion of silica, the 
less readily is the silicate broken up. In order to extract all the lead 
it must first be set free by the use of a stronger basic flux. Thus 
“metallic iron decomposes all fusible lead silicates at a bright red heat, 
provided enough is added to form a singulo-silicate.”” (Hofman). 

Ordinary commercial litharge contains a small amount of silver, 
varying from 0.2 oz. to 1.0 oz. or over per ton. A practically silver 
free variety is made from Missouri lead by giving a zine treatment, as 
for the Parkes process and then cupelling. It is never safe to assume, 
however, that litharge is silver free until it has been proven so by 
assay. Each new lot received should therefore be carefully mixed to 
make it uniform and assayed. See page 89. 


Lead in the granulated form (test lead) is used in the scorification 
assay as a collector of the precious metals and as a flux. When oxid- 
ized by the air of the muffle it becomes a basic flux. Ordinary test 
lead usually contains more or less silver and every new lot should be 
assayed before being used. Lead melts at 326° C. 

Test lead may be made by pouring molten lead just above its freez- 
ing point into a wooden box and shaking it violently in a horizontal 
direction just as it becomes pasty and continuing until it becomes 
solid. The fine material is sifted out, the coarse is re-melted. 


Argols is a reducing agent and basic flux. It is a crude bitartrate 
of potassium obtained from wine barrels. It is one of the best re- 
ducing agents. . 


a 
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5 


Cream of tartar, KHC,H,Osg, is refined bitartrate of potassium. 
Being free from sulphur it is used as a reducing agent in the copper 
assay. Both argols and cream of tartar break up on heating as fol- 
lows :-— 

2KHC,H,O;, + heat = K,0 + 5H.O + 6CO + 2C 
The K;O thus liberated being available as a flux. 


Charcoal, sugar, flour etc. are also reducing agents because of 
the carbon that they contain. Flour is very commonly used in flux 
mixtures and is satisfactory in every respect. 


Sulphides, arsenides, antimonides etc. in ores all have a re- 
ducing effect. 

To determine the reducing power of the different minerals and 
reducing agents see chapter on the crucible assay. 


Iron is a desulphurizing and reducing agent. When heated with 
the sulphides of lead, silver, mercury, bismuth and antimony the 
_ sulphides are decomposed yielding a more or less pure metal and iron 
sulphide. Copper, nickel and cobalt sulphides are partly reduced by 
iron as would be expected by a study of the heat of formation of the 
same. 

It also reduces most of these metals and some others from their 
oxide combinations, as for example :— 

PbO + Fe = Pb + FeO 
The iron oxide formed acts as a basic flux. Iron decomposes all 
fusible lead silicates by replacing the lead, thus:— 
; 2PbO.Si0, + 2Fe = 2FeO.S8i0. + 2Pb 
It should therefore always be used in the lead assay. 
It is used in the form of spikes or nails and sometimes, especially 
in Europe, an iron crucible is employed. 


Potassium nitrate, KNO;, commonly known as niter is a power- 
_ ful oxidizing agent and also a basic flux. It melts at 339° C. and fuses 
without alteration at a low temperature, but ata higher temperature 
breaks up giving off oxygen which oxidizes sulphur and many of the 
' metals, notably lead and copper. 
_ It is used in the fire assay especially to oxidize sulphides, arsenides, 
 antimonides, etc. 
In a charge containing a large excess of soda and litharge the reac- 
tion with pyrite is as follows:-— 
| 6KNO; + 2FeS, + 3Na,CO; = FeO; + 3K.S80O. + 3NaSO. + 
= 3CO. + 3Ne 
4 In this case one gram of niter would oxidize 0.39 grams of pyrite, or 


6 


its oxidizing power would be 4.75 taking the reducing power of pyrite 
in this type of charge as 12.20. 

In a charge containing considerable silica the reaction with pyrite 
is about as follows:— 

1OKNOs + 4FeS, + 2810, = 2FeSi104 + 5K SO, + 3580, + 5Ne 
In this case one gram of niter would oxidize 0.475 grams of pyrite. 
Taking the reducing power of pyrite as 9 in this type of charge, the 
oxidizing power of niter in terms of lead is 4.27. This latter is more 
nearly the figure obtained in practice. As a little oxygen usually 
escapes unused and as the commercial article is never absolutely 
pure, a figure as low as 4.0 is often found about right for practice. 

Niter also reacts with carbon and silica as follows:— 
or 1 gram of niter oxidizes 0.15 grams of carbon, 

2KNOs; + SiO. = KeSiO; + 50 + Ne 
This action begins at about 450° C. 

If finely divided lead is fused with niter, the lead is found to be 
directly oxidized by the niter. Fulton found the oxidizing power of 
niter used in this way to be 2.37. The following reaction shows ap- 
proximately what happens:— 

7Pb + 6KNO; = 7PbO + 3K,0 + 3Ne + 40, 
It should be noted that in this case a considerable portion of the oxygen 
escapes unused. 

Many assayers object to the use of niter because of its oxidizing 
effect on silver. Large amounts of niter cause violent boiling of the 
crucible charge and necessitate careful heating to prevent loss. It 
is found to give less trouble when the crucible is uniformly heated, as 
in themuffle, than when the charge begins to melt first at the bottom, 
a3 in the pot furnace. The student should select an extra large cru- 
cible and carefully watch the fusion when using more than 20 or 30 
grams of niter in any charge. 


Potassium cyanide, KCN, is a powerful reducing and desulphur- 
izing agent. It combines with oxygen forming potassium cyanate, 
thus :— 

PbO + KCN = Pb+ KCNO (reducing action) 
and also with sulphur, forming sulphocyanide, as follows:— 
PbS + KCN = Pb + KSCN 

It is sometimes used in the lead assay and usually in the tin and 
bismuth assays. It is extremely poisonous, and should be handled 
with great care. It fuses at 526° C. 


Salt, (sodium chloride) NaCl, melts at 819° C. and is used as a 





= 
( 


cover to exclude the air, and to wash the sides of the crucible, and 
prevent small particles of lead from adhering thereto. 


It does not enter the slag, but floats on the top of it. 


It is often 


colored by the different metallic oxides of the charge and sometimes 
helps to distinguish assays which have become mixed in pouring. 


Fluorspar, CaF», is occasionally used as a flux. 


It fuses at a high 


temperature, 1400° C., but when melted, it is very fluid and assists in 
liquifying the charge. 


Cryolite, AlNa;F, is not commonly used by assayers, but it is 


sometimes used in melting bullion. 
and has the property of dissolving alumina. 


Name 


Silica 

Glass 

Borax 

Borax glass 

Sodium bicarbonate 
Sodium carbonate 
Potassium carbonate 
Litharge 

Potassium nitrate 
Argols 

Cream or tartar 
Flour 

Charcoal 

Lead 

Iron 

Potassium cyanide 
Salt 

Fluorspar 

Cryolite 


TABLE 1. 





Formula 





S102 
xNasO.yCaO.zS8iO2 
NazgBsO7.10H2O 


KNOs 
KHC4sH4O¢ + C 
KHC4H40¢ 


CaFe 
AlNasF¢ 


ae 
pears: 


Fusion Products. 
Every gold, silver or lead assay fusion, if the charge is properly 
proportioned and manipulated, should show two products, a lead 


button and above it a slag. 








It fuses at a low temperature, 


ASSAY REAGENTS. 





Properties in order of their importance. 








Acid flux 

Acid flux 

Acid flux 

Acid flux 

Basic flux, desulphurizing 

Basic flux, desulphurizing 

Basic flux, desulphurizing 

Basie flux, desulphurizing, oxidizing 
Oxidizing, desulphurizing 
Reducing agent, basic flux 
Reducing agent, basic flux 
Reducing agent 

Reducing agent 

Collecting agent 

Desulphurizing and reducing agent 
Reducing and desulphurizing agent 
Cover and wash 

Neutral flux 

Neutral flux 








Two undesirable products, matte and 


speiss are occasionally also obtained. 
The lead button should be bright, soft and malleable and should 
separate easily from the slag. 
The slag is usually a silicate or borate of the metallic oxides of the 


ore and fluxes used. 
of undecomposed ore. 


It should be homogeneous and free from particles 
A good slag should usually be more or less 


8 


glassy and brittle. When poured, the slag should be thin and fluid 
and free from shots of lead. If too acid, it will be quite viscous and 
stringy, and the last drops will form a thread in pouring. If too 
basic, it will be lumpy and break off short in pouring. When cold, 
the neutral or acid slag is glassy and brittle, the basic one is dull and 
stony looking. 

The slags should never be allowed to get mixed with the fuel, as they 
quickly destroy the furnace lining. 

Matte, is an artificial sulphide of one or more of the metals, formed 


in the dry way. In as aying it is most often encountered in the niter 
fusion of sulphide ores when the charge is too acid. It is found lying 


just above the lead button. It is usually blue gray in color, approach- - 


ing galena in composition and is very brittle. It may be in a layer 
of considerable thickness, or may appear simply as a granular coating 
on the upper surface of the lead button. This matte always carries 
some of the gold and silver and as it is brittle, it is usually broken off 
and lost in the slag, in the cleaning of the lead button. The student 
should examine the lead button as soon as it is broken from the slag, 
and if any matte is found, he may be certain that his charge or furnace 
manipulations are wrong. 


Speiss, is an artificial, metallic arsenide or antimonide formed in 
smelting operations. As obtained in the fire assay, it is usually an 
arsenide of iron approaching the composition of Fe;As. Occasionally 
the iron may be replaced by nickel or cobalt. The antimony speiss 
is very rare. In assaying, speiss is obtained when the iron method is 
used on ores containing arsenic. It is a hard, fairly tough, tin white 
substance found directly on top of the lead and need adhering 
tenaciously to it. 


If only a small amount of arsenic is present in the ore, “ake Speiss. 


will appear as a little button lying on top of the lead, if much arsenic 
is present, the speiss will form a layer entirely covering the lead. It 
carries some gold and silver. If only a gram or so in weight, it may be 
put into the cupel with the lead and will be oxidized there, giving up 


its precious metal values to the lead bath. A large amount of speiss 


is very hard to deal with as it is difficult to scorify. The best way is 
to repeat the assay using some other method. 


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CHAPTER II. 


FURNACES AND ASSAY SUPPLIES. 


Furnaces for assaying may be divided into the two following 


classes :— 


1. Crucible or Pot-Furnaces. These are furnaces used solely 
for melting purposes in which the crucible is in direct contact with the 
fuel or flame and the contents therefore more or less subject to the 
action of the products of combustion. 


2. Muffle-Furnaces. These: are furnaces in which the charge 
to be heated is in a space (the muffle) apart from the fuel or products 
of combustion. The muffle is a semi-cylindrical receptacle of fire-clay 
or other refractory material set horizontally and so arranged that the 
fuel or products of combustion pass around and under it. Thus the 
material to be heated is entirely separated from the products of 
combustion. 

As muffle furnaces may be used for melting purposes as well as for 
scorification and cupellation, many assayers in America use this type 
of furnace exclusively, especially in connection with soft-coal fuel. 
The advantages of muffle furnaces for melting are the greater ease and 


saving of time in charging and pouring, the better control of temper- | 


ature and the better distribution of heat for melting purposes. Cru- 
cibles also seem to stand more heats in a muffle furnace than they will 
in pot furnaces, due no doubt to the slower and more uniform heating. 

Pot-furnaces have the advantage of size, so that for instance in 
dealing with low-grade ores a larger charge and crucible may be used 
than in the ordinary size muffle furnaces. A higher temperature may 
be obtained in pot-furnaces than in muffles and this occasionally is 
an advantage of the pot-furnace. 

The furnaces themselves are made of fire-brick or fire-clay tile and 


“may be set in an iron jacket or surrounded by common red brick. 


Fire-brick is best laid in a mortar made from a mixture of two parts 
ground fire-brick and one part fire-clay. Sometimes a small amount 
of Portland cement is added. In any event the brick and tiles should 
be thoroughly wet previous to applying the mortar. Finally, as little ° 
mortar as possible should be used since the bricks are much harder 
than the solidified mortar. 


10 


Assay furnaces are made to burn practically all kinds of gaseous, 
liquid and solid fuels. Those most commonly used are natural and 


artificial gas, gasoline, kerosene, crude-oil, wood, charcoal, coke, 
bituminous and anthracite coal. | 


Gas is the cleanest, most easily controlled, most efficient in com- 
bustion and except in the case of a natural supply the most expensive 


fuel. A blower is usually required to supply air under a low pressure 
with this method of firing. 


Oil is nearly as clean and as convenient to use as gas, the efficiency” 
of combustion is high and in localities near the oil-fields it may be 
very cheaply obtained. The calorific power of the hydro-carbon 
fuel oils is high, about 50% more than the best coals, which makes 
them particularly suited for use in isolated localities where freight 
charges are high. Gasoline is forced under pressure through a heated 
burner where it is vaporized and the gas injected into the furnace 
carries with it a sufficient supply of air for combustion. Crude-oil 
requires steam or air under pressure to aid in atomizing the oil pre- 
liminary to proper combustion. Gasoline, kerosene and crude 
petroleum all have a heating value of about 21,000 B.T.U. per pound. 


Solid fuels are usually the cheapest and are therefore more ex- 
tensively used than any of the others. In isolated districts where 
coal or coke is not available wood is occasionally used as fuel in assay 
furnaces. For this purpose it should be felled in winter and thor- 
oughly air dried for at least six months or longer according to the 
climate. The air-dried wood will still retain from 20 to 25 per cent 
of water and in this condition has a heating value of about 6000 
B.T.U. per pound. Charcoal is seldom used in this country for assay 
purposes on account of the abundant supply of other fuels. 

Bituminous-coal is the most satisfactory solid fuel for muffle 
furnace firing and coke for pot furnaces. Good soft-coal has a calor- 
ific power of about 14,500 B.T.U. per pound, should be low in sulphur 
and the ash must not be too readily fusible. Coke should be hard and 
strong, low in sulphur and the ash should be infusible at the temper- 
ature of the furnace. That is to say it should be high in silica and 
alumina and low in iron, calcium, magnesium and the alkalies to 
prevent clinkering of the walls of the furnace. 


Gas and Oil vs. Solid Fuel. Gaseous and liquid fuels have many 


advantages over solid fuels for assay purposes, some of which are as 
follows :— 


; 
ne 
: 
: 
3 





ua 


1. The fire is kindled in an instant and the furnace may be quickly 
heated to the desired temperature for work. 

_ 2. The temperature is readily controlled and may be quickly 
varied to suit the requirements of the work. 

3.4A high efficiency of combustion is possible in properly designed 
furnaces and as soon as the work is completed the fuel supply may be 
shut off and fuel consumption stopped. 

4. The avoidance of labor in firing gives the assayer more time for 
other duties. | 

5. The cleanliness in operation due to absence of solid fuel and ash 
is obviously a great advantage in any analytical laboratory. 

On account of the expense however coal is much more generally 
used than either oil or gas. It is easy to make a comparison of the 
costs of any of the fuels by considering the heat units. For instance, 
with soft-coal at $5.00 per ton and gasoline at 15c¢ per gallon we may 
say that one cent invested in soft-coal will buy us 4 x 14,500 = 
58,000 B.T.U. and that the same amount invested in gasoline will 
bring approximately + x 5.4 X 21,000 = 7560 B.T.U. That is to 
say the gasoline is over seven times as expensive as the coal on the 
basis of heat units and for steady running this may be taken to be 
approximately correct. However for a small amount of work a 
gasoline furnace may be cheaper to run even with the cost of fuel as 
above assumed, for the small furnace is quickly heated and as soon as 
the work is completed the oil supply may be shut off and the expense 
stopped, while a coal furnace takes much longer to heat and then must 
be allowed to burn out after the work is completed. 


Coal Furnaces. This type of furnace is used in most of the large 
custom and smelter assay offices in this country. 
The furnace may be built either with a tile or fire-brick lining. The 


- tile lining is more easily set up but whether or not it is as durable as 


a properly constructed fire-brick lining is open to question. The 
outside of the furnace is usually laid up with common hard-burned 
red brick. If the furnace is to be lined with fire-brick several rows 
of headers should be left to hold the lining securely in place. The 
furnace is held together with angle-irons, stays and tie-rods. 

In the furnace as ordinarily constructed the muffles are supported 
by “jamb” bricks projecting from the sides. When these are used 
it is well to leave a hole or lose brick on the outside of the furnace 
to facilitate removing the stubs when these bricks become broken 
off and the ends slagged in. Fulton recommends using long tiles 
here which meet in the center, thus giving better support for the 


12 


muffle. He claims a prolonged life for the muffle with this arrange- 
ment. The writer has found the Scotch Gartcraig brick to outlast 
3 or 4 best American fire-brick for muffle supports. Another method 
of supporting muffles in furnaces of this type is by the use of iron pipes 
or castings extending directly across the furnace and through which 
cooling water is circluated. 

These furnaces occupy a floor space of approximately three and 
one-half by four feet. They are built in a variety of sizes, those 
taking NN, QQ and UU muffles are the sizes; most commonly used. 
The NN muffle is 103 X 19 X 63 inches outside, similarly the QQ 
is 124 x 19 x 72 and the UU is 14 X 19 X 7% inches outside, 

Each NN muffle will hold twelve—-20 gram or eight-—30 gram cru- 
cibles allowing in each case for a row of empty crucibles in front to 
act as warmers, while the QQ muffle will hold fifteen—20 gram .or 
twelve—30 gram crucibles also allowing for a row of empty crucibles in 
front. 

The furnaces are best arranged to be fired from the rear although 
they may be arranged to be fired from the front or sides. The flue 
makes off from near the front of the furnace thus tending to heat the 
muffle uniformly throughout its entire length. It should be from 
one-sixth to one-eighth the grate area. 

The stack for one of the furnaces will need to be at least 20 feet 
high and possibly higher depending largely on the character of the 
coal. It should not be built directly on the furnace but may be placed 
directly over the furnace if supported by arches and cast-iron columns, 
or it may be put to one side of the furnace and in this case will extend 
down to the ground. When the stack is supported independently 
of the furnace it allows the furnace to expand and contract with less 
danger of cracking and also permits of tearing down and rebuilding 
the furnace without interfering with the stack. 

With long-flame coal these furnaces are best fired with a rather 
thin bed of fuel say 8 inches. The sequence of firing will consist 
of first running the slice bar along the entire length of the grate in one 
or two places and lifting up the fire to break up any large cakes and 
thus allow free passage of air through the fire, second to push the 
well coked coal forward with the hoe and third to add 2 or 3 shovels 
of fresh coal near the firing door. As this coal is heated it begins to 
coke and the gas given off passes over the white-hot coal of the fire 
and is there mixed with heated air. This results in a free draft and 
good volume of hot flame. If instead of adding the fresh coal near 
the firing door it is spread all over the fire it will quickly cake and 
tend to smother the fire by shutting off the draft. 





13 


The temperature of the muffle may be regulated at will by man- 
ipulating the draft and firing doors. For instance, after a batch of 
cupels have started the draft may be closed and the firing door opened 
thus admitting cold air above the fire which quickly cools the muffles 
to any required degree. : 


Wood Furnaces. Wood-burning furnaces are made with single 
and double-muffles and are much like the soft-coal furnaces except 
that a larger fire-box and grate are used. Wood is usually sawed 
in half-cord lengths and with dry wood the muffle may be easily 
heated sufficiently for assaying. Hard wood is much to be preferred 
as it does not burn out as rapidly, but almost any kind of dry wood 
may be used. 

The large fire-box and the grate which is set about 8 inches below 
the bottom of the fire-door are the principal distinguishing character- 
istics of a wood burning assay furnace. 


Coke Furnaces. Coke is still used to a considerable extent in 
pot furnaces but for muffle furnace fuel it is fast falling into disuse, 
at least in this country. 

Comparing the coke and the soft-coal muffle furnace the coke furnace 
has the advantage that it can be more quickly heated to a cupelling 
temperature and that it requires less frequent stoking. On the other 
hand it is harder to regulate the temperature, especially to cool it 
off quickly when cupelling, the stoking is harder work and the fuel 
cost per assay is higher in most localities. 

The great advantage of the coke pot-furnace is the very high 


temperature which may be obtained and the fact that even though 


the crucibles boil over or eat through no harm is done to the furnace. 
Coke furnaces especially should be supplied with a good quality of 
fuel. If the ash tends to melt the walls become quickly covered with 
elinkers and are bound to be more or less damaged when these are 
removed. 


Gasoline Furnaces. A gasoline furnace outfit consists of a 
furnace which may be either a muffle, crucible or combination of 
the two, a burner with piping etc. and a gasoline tank with pump. 
The latter for a small assay office consists of an ordinary tinned- 
steel pressure tank equipped with a hand pump, pressure-gauge and 
the necessary piping connections. These range from 2 to 15 gallon 
capacity. 

The burners are usually constructed of special bronze alloys cap- 
able of withstanding oxidation at high temperatures. They consist 


14 


of a filtering chamber for purifying the gasoline, a generating chamber 
where the gasoline is vaporized, a generating pan and valve for the 
initial heating of the burner, a spraying nozzle and valve through 
which the gasoline vapor is injected into the furnace and a,mixing 
chamber where the proper amount of air for combustion is mixed 
with the gas. From the filter the gasoline passes around the interior 
of the burner face (the generating chamber), where it is heated by 
the radiated heat from the furnace and vaporized so that once the fur- 
nace is under way the generating burner may be shut off. Gasoline 
is supplied to the burner under a pressure of from 20 to 50 pounds per 
square inch. 

The great feature to be sought and one of the hardest to attain 
in any gasoline furnace is an even distribution of heat. Another 
feature found wanting in many gas and gasoline furnaces is the poor 
draft through the muffle. Owing to the fact that the pressure in- 
side the furnace is slightly greater than the outside pressure there 
is a great tendency for the products of combustion to work back 
through the hole in the rear of the muffle, thus to a large extent ex- 
cluding the air and unduly prolonging cupellation or scorification. 

In operating a gasoline burner care should be taken to see that 
combustion takes place only in the furnace. All burners have more 
or less tendency to back-fire, that is for the flame to jump back and 
continue in the mixing chamber. If this is allowed to continue the 
burner gets so hot that the metal oxidizes and then it is only a matter 
of a short time before it is entirely destroyed. Every furnace should 
be provided with a shut-off valve between the burner and the gasoline 
tank. When it is desired to shut off the furnace, close this shut-off 
valve letting the burner continue as long as any pressure is left and 
do not ever entirely close the burner valves. The valve stem or 
needle is of steel and the seat is of bronze and owing to the different 
rates of expansion of these metals the valve is injured if these are 
left in close contact when the burner is cooling. This precaution is 
especially to be observed when the burner is provided with the ordin- 
ary needle valve, as when this valve is once enlarged the whole effi- 
ciency of the burner is destroyed. 


Gas Furnaces. Gas furnaces are used in some assay Offices, 
especially where a natural gas supply is available. Where artificial 
gas has to be used this type of furnace proves decidedly expensive 
if used for any considerable amount of work. As the gas is usually 
not under sufficient pressure to carry in its own supply of air for 
combustion, these furnaces are customarily supplied with air from @ 





15 


blower, which adds to the expense and difficulty of the furnace 
operation. 


Crude Oil Furnaces. When a cheap oil supply is available 
crude oil is frequently used as an assay fuel. The furnaces themselves 
are built exactly as for gasoline firing. 

Crude-oil and kerosene cannot be vaporized in the burner as they 
deposit carbon when heated and thus clog the tubes. Consequently 
to insure complete combustion the oil must be thrown into the furnace 
in as fine a state of mechanical subdivision as possible. This is 
accomplished by atomizing the oil with a jet of steam or air. The 
steam may be from an outside source or may be generated in the face 


of the burner from the heat reflected: from the furnace. The Marvel 


burner made by the Braun Corporation of Los Angeles uses the latter 
method of supplying steam. This burner requires to be heated by | 
a torch or alcohol lamp to start and the necessary outfit includes two 
pressure tanks one for oil and one for water. Oil of any gravity up 
to 60°, Beaume may be used with this burner. 


Furnace Repairs. 


Fire-clay usually forms the basis of mortars used in furnace con- 
struction and repairs, as lime mortar and hydraulic cement are not 
suited for use with masonary exposed to high temperatures. Fire- 
clay is a clay containing only very small amounts of iron, lime, mag- 
nesia and the alkali oxides. It forms a more or less plastic and sticky 
mortar and on heating loses its moisture and plasticity and the mortar 


hardens. 


All clays shrink more or less on drying and burning and to prevent 
this as far as possible in the mortar as well as to make it strong a 
certain amount of crushed fire-brick or sand should be added. 
Crushed fire-brick is better than sand owing to its porous and ir- 


regular shaped. grains as these give a better mixture with the clay 


and a stronger cement. 

A good mortar for general use around assay furnaces is made with 
a mixture of two parts ground fire-brick through 12 mesh and one 
part fire-clay.. A small amount of Portland cement or molding clay, 
say not over one-third part, will make the mixture adhere better and 
the mortar will be harder when set. For work at very high tempera- 


a tures the Portland cement must be omitted as it acts as a flux for the 





other materials and causes the whole to melt. 
All mortars should be made up dry and thoroughly mixed before 


4 - the required amount of water is added. The water should be thor- 


16 


oughly mixed in and the mortar should be sticky and of the right 
consistency. It is well to mix the mortar several hours before using. 
When laying bricks or making repairs about a furnace the bricks 
and-brick work should be thoroughly wet before applying the mortar 
as otherwise the bricks absorb so much water that the mortar does not 
form a good bond with them. 

In laying fire-bricks as little mortar as possible should be used as 
the bricks are always harder than even the best of mortar. The mor- 
tar should be made to fill every crevice. The best way to attain this 
is to put an extra amount of fairly thin mortar on the wet brick and 
then drive or force it firmly into place, allowing the excess mortar to 
squeeze out. 

The ash from many coals is quite readily fusible and results in the 
formation of clinkers and accretions on the sides of the furnace, 
especially just above the grate. When the furnace is cold these 
adhere very tenaciously to the walls of the furnace and in breaking 
them off, pieces of the brick are removed with them. To remove these 
accretions with the least damage to the furnace they should be cut 
off with a chisel bar just after a hot fire has been drawn. 

In putting in a new muffle, first remove the old one with the mortar 
that held it, also any clinkers which would interfere with the working 
of the furnace. Patch the lining of the furnace if it requires it and 
see that the bricks or other supports for the muffle are in place and 
in good condition. After trying the muffle to see that it rests properly 
on the supports, remove it, sponge over the brick work where the mor- 
tar is to come in contact with it, place some rather thick mortar on 
each of the supports and replace the muffle. See that it rests evenly 
on the different supports and on the front wall of the furnace. The 
muffle should be level horizontally and slope slightly toward the 
front end. Fill up the space between the muffle and the front wall 
of the furnace with some rather thick mortar, working from both 
inside and outside of the furnace. This outside joint should be finished 
up neatly with the aid of a trowel. It is best to allow the furnace to 
dry for a day or two if possible, but if necessary it may be used as 
soon as finished by heating up slowly. 

For patching the linings of furnaces use the mixture recommended 
for general use or try the following which is recommended by Lodge. 
Fire-brick through 12 mesh 7 parts, Portland cement 2 parts, fire- 
clay 1 part. Put this on as dry as possible and it will make a patch 
almost as hard as the original brick. 

Cracked and broken muffles may be made to last much longer if 
patched with some of the following mixtures. 





ee ee a ea ee ey 





17 


~ When the bottom is almost gone Lodge recommends a mixture of 
2 parts Portland cement, 1 part ground fire-brick, 4 to $ part fire- 
clay. For patching holes he recommends a mixture of glass, sand and 
clay to which a little litharge has been added. After heating this 
becomes as hard as the muffle. A mixture of short fibered asbestos 


and silicate of soda is also recommended. 


Metallurgical Clay Goods. 


Under the caption metallurgical clay goods are included muffles, 
crucibles, scorifiers, roasting dishes, annealing cups ete. These 
embrace many of the most important utensils of the assayer and 
upon their good properties much of his success depends. Fire-clay 
is the only material which answers the double purpose of satisfactory 
service and inexpensive construction. Refractory clay or fire-clay 


as it is commonly called is a clay which will stand exposure to a high 


temperature without melting or becoming in a sensible degree soft or 
plastic. 

All clays contract both upon drying and burning and this leads to 
more or less warping and cracking of the finished product. ‘To pre- 
vent this shrinkage as far as possible and also to add strength to the 
finished article it is customary to add a certain amount of sand or 
well burned clay to the mixture. Burned clay is usually preferred 
to sand for this purpose, not only because its rough porous grains 
give a better bond with the fire-clay and make a stronger cement, but 


it also makes an article which is less readily corroded by assay slags 


and fusion products. The intermixture of coarse grains of burned 
clay helps also in that it makes a product better able to withstand 
sudden changes in temperature. 

The exact proportions of raw and burned clay used by any manu- 
facturer are carefully guarded trade secrets and depend of course 
very much on the clay used as well as upon the article to be manu- 
factured. The larger the article the more is the care which must be 
taken to prevent warping and cracking. Usually however, the pro- 
portion of raw to burned clay will lie between the limits of one to one 
and one to two. 


Muffles. Muffles may be made of a variety of materials but for 
assay purposes fire-clay muffles are used exclusively. They are 
made in a great variety of sizes and shapes. 

Muffles as well as other fire-clay ware should be stored in a warm, 
dry place and should be heated and cooled slowly and uniformly for 


18 


maximum service. The life of a muffle is also much influenced by 
the manner of supporting. 


Crucibles. Assay crucibles are made either of a mixture of raw 
and burned clay or of a mixture of sand and clay, the first being known 
as clay or fluxing crucibles and the second as sand crucibles. The 
raw clay is finely ground, mixed with the right proportion of coarser 
particles of sand or burned clay and water and the whole well kneaded 
and compressed in molds of the proper shape. 


Good crucibles should have the following properties:— 
. Ability to withstand a high temperature without softening. 
. Strength to stand handling and shipping without breaking. 
. Ability to stand sudden changes of temperature without cracking. 
. Ability to withstand the chemical action of the substances fused in 
them. 
5. Impermeability to the substances fused in them and to the products 
of combustion. 


me Wh Ke 


Of course it is impossible to get any one crucible which will possess 
all of the above good properties to a high degree. For instance if a 
crucible is to be made as nearly impermeable as possible it will be 
made of very fine grained material and tightly compressed. Such 
a crucible however will not stand handling or sudden changes of tem- 
perature as well as one made with a skeleton of coarser material. 
Furthermore the manner and temperature of burning has much to 
do with the way that crucibles will stand handling and shipping. 
A fairly hard burned crucible will be stronger and less likely to be 
broken in handling but on the other hand it will not stand sudden 
changes of temperature as well as a soft-burned crucible. Crucibles 
made of elay containing little uncombined silica and of burned elay 
of the same nature will stand a high temperature and chemical cor- 
rosion much better than those made of sand and clay or of elay eon- 
taining much free silica. 

Crucibles are tested for resistance to chemical corrosion by actual 
service and also by fusing litharge in them and noting the time it 
takes to eat through. To make a test of this sort which is of any value 
care must be taken to see that the temperature, the quantity of 
litharge and all other conditions are the same for the crucibles being 
tested. A crucible may be tested for its permeability to liquids by 
filling it with water and noting the time it takes before it becomes 
moist on the outside. 

Crucibles come in a great variety of shapes and sizes. Those most 





19 


commonly used for assaying may be classified into two groups as — 
follows :— 


Pot Furnace Crucibles. These are comparatively slim, heavy 
walled crucibles with practically no limit as to height. The base is 
small so that they may be forced down into the fuel and for this reason 
they are easily tipped over and are not suitable for muffle work. 
The sizes most used are the E, F, G, H, J and K. Crucibles of the 
same designation made by different manufacturers vary considerably 
in capacity. The approximate capacity of some of the pot-furnace 
crucibles is shown in the following table:— 


TABLE II. CAPACITIES OF POT FURNACE CRUCIBLES. 




















Crucible designation E F G H I J kK 
1 Battersea 180cc | 210ce | 300ce | 420cec - 600cce | 750ce 
2 Denver 180ce | 240ce | 400cec —% 530ce 685ee | 950ce 











\ Made by the Morgan Crucible Co. London, England. 
2 Made by the Denver Fire Clay Co. Denver, Colorado. 


Muffle Crucibles. These are made with a broader base so that 
they may stand securely on the floor of the muffle and are usually 
not more than four inches high. Muffle crucibles are designated by 
gram capacity, the 10, 15, 20 and 30 gram sizes being most frequently 
used. The intention of the system is that the numbers indicated 
the grams of ore charge which the crucibles will take. They are 
usually generously proportioned so that often an assay ton of ore 
(29.166 grams) may be treated in a 20 gram crucible. 

The approximate capacity of the more important muffle crucibles is 
shown in the. following table:— 


TABLE III. CAPACITY OF MUFFLE FURNACE CRUCIBLES. 


Crucible Designation Sem.) 10 ems) 12. gm. | 15 gm. | 20 gm. | 30 gm. 


2) | ee, ee en 


Denver 70cec | 100ce | 140ce | 160cce | 190ce | 260ce 
Battersea 70ce. | 100ce - 135ee | 190ce | 260ec 











Scorifiers. These are shallow fire-clay dishes used in the scorifi- 
cation assay of gold and silver ores. They should be smooth on the 
inside, dense and impermeable to lead and slag and should be composed 
so as to withstand as much as possible the corrosive action of litharge. 
Scorifiers are designated by their outside diameters. Of the large 


20 


number of sizes made the following are the most commonly used: 
24/7, 24/", 22", 3’’, 34’".. The Bartlett scorifier is shallower than the 
regular one and was designed for the treatment of heavy sulphide 
ores containing considerable metallic impurities. Scorifiers particu- 
larly should be made of clay containing a minimum of uncombined 
silica, as the scorifier slags are usually very basic. Particularly 
when they contain copper they attack the silica of a scorifier with 
avidity and one with a siliceous skeleton may become perforated 
and allow its contents to escape onto the floor of the muffle, thus 
spoiling the assay and injuring the muffle. 


2. + oe ee ee 


yore ae 


Ee ee Ee ae ee ee Le ee me ee 





CHAPTER III. 
SAMPLING. 


Definition. A sample is a small amount which contains all the 
components in the same proportions as they occur in the original lot. 
The object of sampling an ore is to obtain for chemical or mechan- 
ical tests a small amount which shall contain all the minerals in the 
same proportion as they occur in the original lot. In the subsequent- 
discussion the word “sample” will be taken to mean that fraction 


which is taken to represent the whole, whether or not it does so. The 


compound words correct-sample, representative-sample, true-sample, 
will be used to represent the ideal conditions. 

In the intelligent operation of a mine or metallurgical plant, it 1s 
necessary to sample and assay continually. In most mines, the differ- 
ent faces of ore are sampled every week, sometimes every day. In 
concentrating plants, it is customary to sample the products of every 
machine to ascertain whether the machine is doing the work expected 
of it. In smelters, every lot of ore, as well as fluxes and fuels, have 
to be sampled and assayed in order to calculate a charge which will 
run properly in the furnace. The slag, flue dust and metallic products 
must also be sampled and assayed in order to maintain control of 
the operations. In lixiviation plants, the ore and tailings as well 
as the solutions must be sampled in order to control and check the 
daily work of the plant. In fact, careful sampling and assaying can 
not be disregarded, and is becoming more and more important every 
day as the grade of ore decreases and the margin of profit becomes 
less. 

The assayer will usually have the major part of the sampling done 
for him, but he is expected to know how to do it when called upon. 
He will usually have only to prepare the final sample, but will oc- 
casionally receive lots of 10 to 100 or more pounds to assay in which 
case he will have do to his own sampling. The following discussion 
will deal principally with the assay laboratory problems of sampling 
and the questions of mine and mill methods will be omitted. 


22 


Labelling Samples. Every lot of ore coming into an assay office, 
laboratory, custom mill or smelter should be given a lot number which 
should never be repeated, and should be immediately labelled with 
this number. A record book should be kept for this purpose, should 
show the number of the sample, date of receipt, name of mine, com-~ 
pany or individual from whom received, the gross and net weight, as 
well as notes on the general mineral character, ete., ete. 


Moisture Sample. Assays and chemical determinations are 
always made on dry samples and the value of a lot of ore is always 
figured on the moisture free basis. Exeept in cases when the entire 
lot may be dried, it is necessary to take a sample from which to de- 
termine the moisture. This sample must be taken as quickly as 
possible after the ore is weighed. If the ore may be quickly crushed 
and sampled to a small amount of 12 or 14 mesh material, the mois- 
ture sample should be taken from this and put in a closely covered 
pail or box. Duplicate samples of one or more kilograms of this are 
weighed out into a porcelain or enamelled iron dish and dried at 110° C. 
The loss of weight being called moisture. As the sample is handled 
more than the reject, it loses some moist ire enroute and a constant 
should be added to compensate for this difference. Brunton’ finds 
10 per cent in summer and 7 per cent in winter a fair average figure. 
For instance, if the sample showed 5 per cent moisture for a lot of 
ore shipped during the summer months a fair figure for the actual 
moisture content would be 5.5 per cent. | 


Operations. 


Ore sampling may usually be considered to consist of three distinct 
operations repeated as many times as necessary. These operations 
are Ist, crushing; 2nd, mixing; 3rd, cutting. After the cutting we 
have a sample and a reject. The sample may be further reduced by 
a repetition of the three operations until it has reached the desired 
bulk. 

The whole science of ore sampling depends primarily on a eorrect 
knowledge of the proper relation between the maximum size of the 
ore particles and the weight of the sample taken. The problem to 
be solved in each case is something as follows:—having crushed a 
particular ore to a certain size (say 10 mesh), how small a sample is 
it safe to take from this and still keep within the limit of error allow- 

'T. A.M. EH, XL) pg: 5672 (1909) 


. 
; 
“4 
4 





able? It is necessary to know thé ore, the limit of error allowable, 


and the mathematical principles involved. 

Sampling is classed as hand sampling when the mixing and cutting 
down is done by men with shovels and as machine sampling when 
done by some form of automatic machine. 


Crushing. All the of ore, unless already fine enough, is broken 
or crushed to pass some limiting size screen. This size depends upon 
the value of the ore and other factors to be considered later. The 
finer the pulp is crushéd the more uniform in size are the particles 
and more thorough mixing and better sampling is possible. If the 
ore is to be smelted, most of it should be left in the coarse state as 
fine ore is undesirable. If it is to be roasted or leached, on the other 
hand, fine ore is not objectionable, and the first crushing may be 
carried further. Asa rule, however, the aim is to minimize the crush- 
ing, thus saving in cost and keeping down the dust. 

Machines for crushing should be rapid in action and capable of 
easy cleaning. Jaw breakers and rolls fulfil these requirements, ball 
mills and pebble mills do not. 


Mixing. This step in the process of sampling is often omitted or 
allowed to be taken care of itself. It is a necessary forerunner of 
quartering and channeling, but is usually omitted before the other 
methods of cutting. Especially in the handling of small lots of ore 


in the laboratory, it is best to be over careful in this particular rather 


than the reverse, and, as it adds but little labor, to give each lot of 
crushed ore a thorough mixing before cutting. 

The four following methods are used in assay office sampling, some 
being better suited for large lots and some for small lots. 


1. Coning. The sample is shovelled into a conical pile, each 
shovelful being thrown upon the apex of the cone so that it will run 
down evenly all around. In mixing a large lot of ore by coning, it 
is first dumped in a circle and then coned by one or more men who walk 
slowly around between the cone and the circle of ore. The best 
results are obtained by coning around a rod, as by this means the 
center of the cone is kept in a vertical line. Coning does not thor- 
oughly mix an oré, but rather sorts it into fine material which les 
near the center and coarser which rolls down the sides of the cone. 
If the ore is practically uniform in size, and specific gravity, the mix- 
ing may be more thorough. A slight dampening of the ore is said to 
allow of better mixing by coning. The floor for this and other hand 


24 


sampling operations should be smooth and free from cracks which - 
would make good cleaning difficult or impossible. A floor made from 
sheet iron or steel plates is preferable. 


2. Rolling. For lots of 200 pounds or less the method of mixing 
whereby the ore is rolled on canvas, rubber sheeting or paper is often 
used. When the ore particles are fairly uniform in size and specific 
gravity, this method is satisfactory, but for ordinary ores in the coarse 
state, it should be avoided. For ore crushed so fine that it has little 
or no tendency to stratify, as for example the assay pulp ground to 
100 or 120 mesh, the method has been found satisfactory when the 
operation is properly performed. ‘This method is almost universally 
used by assayers for mixing the final lot of pulverized ore just before 
taking out the assay portion. — 


3. Pouring. For small samples the method of pouring from one 
pan into another is sometimes employed, especially as a preliminary 
to rife cutting. Like the two above, it is imperfect when performed 
on ordinary coarse and fine ore mixed. 


4. Sifting. For mixing small lots of ore or fluxes the method of 
sifting is particularly good. The apertures in the sieve should be 
two or three times as large as the largest particles. The ore should 
be placed on the sieve a little at a time and allowed to fall undisturbed 
into a flat receiving pan until all the ore has passed the sieve. Two 
or three siftings are equivalent to 100 rollings. Sifting has the further 
advantage over all the other methods that all lumps are broken up 
and the ore composing them distributed. 


Cutting. The final step in the sequence of sampling operations 
consists in taking out a fraction of the whole, say one-quarter or one- 
half, in some systematic impartial manner. The part taken out is 
called the sample and the operation of taking it is the cutting. 

The four following methods of hand cutting are used considerably, 
but some of them are giving way to machine sampling methods. 


1. Fractional Selection. This is a rough starting method suited 
only to large lots of low grade or fairly uniform ore. When the ore 
is being taken away from the crusher or shovelled out of cars as the 
case may be, every second, third, fifth, tenth shovelful, depending 
on the value and uniformity of the ore is taken and placed in a sep- 
arate pile which is afterwards cut down by some of the later described 


| 
: 
: 
| 
: 





25 


methods. When shovelling the ore, care must be taken that each 
shovelful is taken from the floor. In case the ore contains lumps too 
large for the shovel, they should be broken and put back on the pile. 
The method is open to the serious objection that it is a very simple 
matter for a prejudiced party to make the sample either higher or 
lower in grade than the average by selection of his shovelful samples. 


_ 2. Channeling. This consists in spreading out the crushed and 
mixed ore in a flat layer a few inches thick and then taking the sample 
out in parallel grooves or channels across the pile’ Often two sets 
of channels are made one set at right angles to the other. Channeling 
is a slow method, requiring much labor and floor space, and owing to 
the coarse pieces which fall into the channels from the sides it is in- 
accurate. ‘The method is fast falling into disuse. 


3. Quartering. This is the method of cutting which accompanies 
coning. It presupposes a thorough mixing by coning, as the two 


always go together. 


When the cone is completed, it is worked down into the form of a 
flat truncated cone by men who walk around and around drawing 
their shovels from center to perifery, or by starting at the apex and 
working the shovel up and down in the path of a spiral. The point 
to be observed here is not to disturb the radial distribution of the 
coarse and fine ore. After flattening, the cone is divided into four 
90 degrees sectors or quarters by means of a sharp edged board, or 
better by a steel bladed quarterer. These quarters should of course 


radiate from the position of the center of the original cone. Two 


opposite quarters are taken out and rejected and the two others are 
then taken for the sample. This sample may be again mixed by 
coning and quartered, or crushed, coned and quartered as the case 
may require. 

When properly carried out the method may be made to yield fairly 
accurate results, but at best it is a slow and tedious process, and re- 
quires the most conscientious work on the part of the laborers to 
insure correct results. It is open to the objection that it affords 
opportunity for manipulation of the sample by dishonest operators. 

Coning and quartering is the old Cornish method of ore sampling 
and was almost universally used 30 years ago. It is still somewhat 
used as a finishing method at sampling works and by engineers in 
the field where no machinery is available. 


4. Riffle Cutting. Riffle cutting or splitting is the most accurate 
laboratory method available. The riffle, splitter or split-shovel 


26 


consists of a number of parallel troughs with open spaces between 
them, the spaces being usually the same width as the troughs. These 
troughs are rigidly fastened and the whole is made into the form of a 
shovel, called a split shovel. The ore is taken up on a flat shovel or 
special pan and spread over the troughs, care being taken to prevent 
heaping the ore above the troughs. Either the ore which falls in 
the troughs or that which falls between them may be taken as the 
sample. The cutting may be repeated as many times as is deemed 
desirable. For the best results in cutting any sample of ore by this 
method, care should be taken to have only a thin stream of ore fall- 
ing from the pouring pan and to move this pouring pan back and 
forth over the split shovel in a horizontal direction perpendicular 
to the riffles, so that every part of the stream of ore is being directed 
alternately and rapidly first into the sample and then into the reject. 
The more irregular in size, specific gravity and value are the minerals, 
the greater the care which should be taken in this particular. The 
sample should be mixed before re-cutting. 

A modification of the riffle or split-shovel known as the Jones 
Sampler. or simply as a ‘“‘splitter’’ has recently come into use. It is 
a riffle sampler in which the bottoms of ‘he riffles are steeply inclined, 
first in one direction and then in the o.her. The ore is spread over 
the riffles in the Jones Sampler exactly as over the split-shovel. The 
ore 1s caught in two pans placed underneath the splitter. 

Riffle cutting is the most rapid method of hand sampling and is 
also the most accurate. It is used as a finishing method in most 
modern sampling works. One objection to the Jones Sampler and 
other similar models is the considerable amount of fine ore-dust which 
may be lost due to the greater length of fall of the ore before coming 
to rest. One way to obviate this would be to slightly moisten the 
thoroughly mixed ore before cutting. 

In selecting a split-shovel or riffle cutter for any particular sampling 
operation, care should be taken that the distance between the riffles 
be at least three times the diameter of the maximum particle of ore. 
It is found that a slight bridging action may occur if this precaution 
is not observed. 


Machine Cutting. A large number of machines have been de- 
vised to take the place of the slow laborious methods of hand sampling. 
They all depend on taking the sample from a stream of falling ore. 
All these devices for sampling fall either under the head of continuous 
or intermittent samplers. The continuous samplers take part of the 
stream all the time by placing a partition in the falling stream of ore 








27 


to separate sample from reject. The intermittent samplers, as the 
name implies, deflect the entire stream at intervals to make the 
sample. 

The continuous method of sampling is open to the objection that 
it is impossible to get a stream of falling ore containing coarse and 
fine particles which is uniform across its entire section. Therefore, 
any continuously taken sample (except possibly one-half of the stream) 
will be either richer or poorer than the average. Because of these 
conditions this type of sampler will not give uniformly reliable re- 
sults, and is now but little used. 

The intermittent method of sampling gives better results. The 
machine should be so designed that it takes equal portions all across 
the stream at regular intervals. While it is not possible to produce a 
stream of ore which is uniform in value throughout its entire length, 
yet by taking a large number of small samples entirely across the 
stream the average thus obtained gives a good representative sample 
of the entire lot. It is essential that the percentage of sample taken 
from all parts of the delivery pipe be the same, in other words that the 
vertical sample section, taken in a direction parallel to the motion of 
the intake-spout should be a rhomboid. 

Three machines of this type have come into general use; these 
are the Brunton, the Vezin and the Charles Snyder. The Brunton 
oscillates in a vertical plane through an arc of 120 degrees and by a 
change of gears any proportion of the stream from 5 per cent to 20 
per cent may be taken. The Vezin and Charles Snyder machines 


have sector shaped sample cutters radiating from a shaft around which 


they revolve. 


Hand and Machine Sampling Compared. In comparing hand 
and machine sampling it may be said that machine sampling is gen- 
erally cheaper and with a properly designed machine it is more ac- 
curate than coning or fractional selection. Perhaps the most im- 
portant advantage of all is that being strictly mechanical in operation 
it affords less opportunity for manipulation of the sample. 


Precaution to be Observed. Besides the danger of ‘‘salting”’ 
from crushing machines, elevators, sampling machines etc. special 
attention must be paid to the disposition of the fine ore dust. Asa 
rule the rich minerals in the ore are more brittle than the gangue 
with the result that the ore dust is far higher in grade than the average 
of the ore. Whence is seen the necessity of preserving all of the ore 
dust and of taking pains to see that the sample contains its proper 
proportion of the same J 


28 


Theoretical Considerations. The most certain method of 
obtaining a representative sample of a lot of ore would be to crush 
the whole to 100, 120 mesh or finer, mix it thoroughly and then cut 
down by one of the methods just described. This method can be 
followed for small amounts of a pound or so, but in the ease of large 
lots, it would entail too much labor and would usually unfit the ore 
for future treatment. The method generally adopted is a compromise 
and consists in crushing the whole lot to a certain predetermined 
maximum size and then taking out a certain fraction as a sample. 
This sample is again crushed to a smaller size and cut down as 
before and this process repeated until finally the assay sample is 
obtained. 

The care which is required in sampling as well as the size to which 
a lot of ore or other material must be crushed before a sample is taken 
depends upon the value and uniformity of composition of the material. 
The more uniform it is, the smaller may be the sample taken after crush- 
ing to any particular size. For instance, if we have a solid piece of 
galena containing silver uniformly distributed as an isomorphous silver 
sulphide, we can break off a piece anywhere, crush it and have for 
assay a lot of ore which is truly a sample of the piece. If, however, 
our specimen is not solid galena, but is made up of galena and lime- 
stone, the silver value still being contained in the galena, we will 
have to crush the whole lot to a uniformly fine size before taking out 
a fractional part for a sample. Furthermore it will readily be seen 
that the greater the difference in the grade of the different minerals 
in the ore, the finer must a lot of ore be crushed before a certain sized 
sample should be taken from it. 

Since ores are never perfectly uniform in composition a certain 
amount of crushing is evidently necessary in every case. To determine 
the amount of crushing we must first consider the commercial side 
of the question, that 1s we must determine how far it will pay to go 
with the process. Evidently a mistake of 1 per cent in the iron 
contents of a car load of iron ore worth say $3.00 a ton is less serious 
than the same percentage error in the copper contents of a car of cop- 
per ore worth say $50.00 a ton. ‘Therefore it may be seen that it 
will pay the seller or buyer of the copper ore to go to more pains and 
expense in the sampling of the ore than if he were dealing with the 
less expensive iron ore. | 


Varying Relation of Size of Sample to Maximum Particle. 
The variation of any portion of a lot of ore from the average com- 
position of the whole is due to the excess or deficit of one or more 





4 
: 
; 
) 
; 
| 
‘ 
; 
7 
| : 


29 


particles. The effect upon the results will be greatest when the piece 
or pieces which are in excess or deficit are of the largest size, greatest 
specific gravity and greatest variation in quality from the average. 
Disregarding for the moment the last two of these factors and sup- 
posing the ore particles to be approximately uniform in size it is evi- 
dent that the sample must contain enough particles so that one ad- 
ditional particle of the richest mineral would practically cause no 
variation in the value. This means that the sample of the ordinary 
ore must contain a very large number of particles perhaps 500,000. 
Having determined how many particles of the ore it is necessary 
to include in the sample, and assuming the different minerals to be 
entirely detached from one another, it would be fair to take such a 
weight of ore after each reduction as would contain this established 
number of particles. Or as the weight of a lump is proportional to 
the cube of its diameter we may state the rule as follows:—Make the 
weight taken for the sample proportional to the cube of the diameter 


- of the largest particle of the ore. 


In the ordinary ore, however, the different minerals are not entirely 
detached from each other, but approach more and more to this con- 
dition as the size of the ore is reduced. Hence a fixed number of the 
particles of the fine ore is less likely to be a true average of the whole 
than the same number of pieces of the lump ore before it was broken. 
Therefore as the size of the ore is reduced a larger and larger number 
of particles should be taken for the sample. To conform to this con- 
dition of affairs the following rule was proposed by Professor R. H. 
Richards: ‘‘For any given ore the weight taken for a sample should 
be proportional to the square of the diameter of the largest 
particle.”’ 

The accompanying table embodies this rule and is based on figures 
taken from the practice of several careful managers. It was arranged 
and is now published with the permission of Professor Richards. 

The first column shows the safe weight in pounds for a sample of 
ore of any of the six grades shown and for sizes as indicated:in the 
respective columns. Column 1 applies to iron ores, column 2 to low 
grade lead, zinc and copper ores and even to low grade pyritic gold 
ores, without native gold, where the pyrite is evenly distributed 
through the ore. Columns 3 and 4 apply to ores in which the valuable 
minerals are less uniformly distributed. Columns 5 and 6 apply 
to ore containing fine particles of native gold or silver, also to telluride 
and other “spotty ores.” 


30 


TABLE IV. WEIGHTS TO BE TAKEN IN SAMPLING ORES, 


. a - sak ed 
ees 














1 2 | 3 4 5 6 
etch te Of Diameter us Largest Particles Millimeters: 
Sample “i ; 
Very Low Low ‘ Very 
cons Grade or | Grade or Medi 0 ae Rich and 
Very Uni-} Uniform papetarite ae i Spotted 
form Ores. Ores. ek Ores. 
20,000.000 207.00 | 114.00 76.20 50.80 31.60 5.40 
10,000.000 147.00 80.30 53.90 35.90 92.40 3.80 
5,000.000 107.00 | 56.80 38.10 25.40 15.80 2.70 
2,000.000 65.60 | 35.90 24.10 16.10 | 10.00 170 
1,000.000 | 46.40 25.40 17.00 11.40 7.10 1.20 
. 500.000 32.80 | 18.00 12.00 8.00 5.00 85 
200.000 20.70 | 11.40 7.60 5.10 320 54 
100.000 14.70 8.00 5.40 3.60 2.20 38 
50.000 10.70 5.70 3.80 2.50 1.60 27 
20.000 | 6.60 3.60 2.40 1.60 1.00 17 
10.000 4.60 2.50 1.70 1.10 71 12 
5.000 | 3.30 1.80 1.20 80 50 
2.000 | 2.10 1.10 76 5l a2 
1.000 1.50 80 54. 36 22 
500 1.00 57 38 25 16 
200 | 66 36 24 16 10 
100 A6 25 nz 1 
050 3G 18 12 
.020 21 11 | 
010 15 
005 | 10 | | 














It should be remembered that the above mentioned rules for sam- 
pling will not hold for ore containing large pieces of malleable minerals 
such as native gold, silver, silver sulphide, chloride etc. These roll 
out and do not crush and must be treated by special methods. See 
“Sampling Ores Containing Malleable Minerals.” 

In using the table, it is not necessary to crush successively to all 
of the sizes shown in any of the columns. The ore may be crushed 
to any fineness convenient and then a weight of sample corresponding 
to that shown in the table may be taken. In sampling mill practice 
it is customary to reduce the diameter of the coarsest particles one- 
half at each stage or crushing, thus reducing the volume to one-eighth 
or 12.5 per cent. It is also customary in practice to take a 20 per 
cent sample at each stage, consequently the ratio between the weight 
of sample and size of maximum particle is constantly increasing 
throughout the sampling process, thereby meeting theoretical condi- 
tions previously discussed. 





ol 


Relation of Size of Sample to Grade of Ore and Effect of 
Specific Gravity of Richest Mineral. Although it had long been 
appreciated that the size of the sample would have to be greater as 
the richness of the ore increased, it remained for Reed’ to develop 
a formula by which the proper ratio between these could be scientifi- 
eally maintained. 

D = diameter of largest pieces in inches. 

P = quantity of the lot in Troy ounces. 

f = number of parts into which P is to be divided before one part 

is taken as a sample. 
= percentage of silver or gold in the richest specimens in the lot. 
= specific gravity of the richest minerals. 
average grade of ore (ounces per ton). 
= number of pieces of D size and k value that can be in excess 
or deficit in the portion chosen as sample. 
largest allowable percentage of error. 


So By 
II 


l 


Before taking a sample of ; Troy oz. we must crush the lot so that 


D = 05 PL mPl 
sk(f-la 


- Taking for general purposes s = 7, | = 1 and a value of a of 1.6, 
the following table is given for different grades: 


Medium m = 50 el 

High grade m = 75 Kea 

Very rich m = 500 Ko=230 
TABLE V. 











a ay 





Value of D in inches. 
Sample reduced from 





Medium High Grade | Very Rich 








100 to 10 tons 


5.28 2.96 2.53 
3 10to ) 1 ton 2.46 1.38 Be 
2000 to 200 lbs. 1.14 0.6 0.56 
200 to 5 lbs. 0.3 0.18 0.16 
5 lbs. to 10 assay tons 0.0 





This quantity assumed for a appears to be very small and in the 
opinion of the writer should be larger, which would have the effect of 
reducing the values for D shown in the above table. 

Brunton’ derived a formula similar to Reed’s but more convenient 
to use. 


1School of Mines Quarterly VI. page 351 (1885) 
2T. A.I. M.E. XXV. p. 826 (1895) 


32 


W = weight of sample in pounds. 
k = grade of richest mineral in ounces per ton. 
c = average grade of ore in ounces per ton. 
s = specific gravity of richest mineral. 
n = number of maximum sized particles of richest mineral in éx¢éss 
or deficit in sample. 
f = a factor expressing the ratio of the actual weight of the largest 


particle of richest mineral which will pass a screen of a given 
size to the weight of the largest cube of the same mineral which 
will pass the screen. 
p = allowable percentage error in sample. 
D = diameter in inches of the holes in the screen, or other normal 
diameter to which the ore is crushed. 
From these he finds 


D = 65 


Making p, the allowable percentage error, = 1, the formula be- 
comes: 

7 = ee 

D = .65 Noe 

fsn(k—c) 

To determine a value to use for n Brunton made a number of assays 
on two different lots of high grade ore crushed to pass a certain limit- 
ing screen. The average deviation from the mean = p was substi- 
tuted in the formula and results of 2.64 and 3.14 respectively were 
found for n. Assuming that 3 is a safe value for n and cubing each 
side we find. 





D? an We 
~ 10.8 fs(k-c) 
or 
ae 10.8fsD*(k—e) 





x 
from which may be found the safe weight in pounds for 4 sample 
of any ore whose largest particle is D inches. Taking four examples 
using as the richest minerals pyrite, galena, native silver and native 
gold and assuming different values for D, k, ¢ and f the following 
table was made after the style of the table first shown in Hofman’s 
Metallurgy of Lead. The values for f used for the fine sizes were 
those determined by Brunton’s experiments, i.e. 4 for pyrite and — 
galena and 6 for native silver and gold. This value of f is reduced 
gradually until for 1 inch diameter, it is made equal to one, this 








a ee eee ee eee 


33 


variation therefore tending to compensate for the greater uniformity 
of value of the particles as they become larger. 

This following table is probably the best and certainly the most 
conservative of all. A good deal of intelligent discrimination may 
often be used however and mere formula can never be made to cover 
all possible contingencies. For instance in sampling an ore in which 
the valuable mineral is finely and uniformly disseminated throughout 
the gangue, a much smaller sample than that given in the table may 
be taken for the coarse sizes, although for the fine sizes the full quanti- 
ties shown in the table should be taken. Another ore, with perhaps 
the same ratio of value of the richest mineral to average grade, having 
the rich mineral in larger crystals or masses will have to be sampled 





















TABLE VI. WEIGHTS TO BE TAKEN IN SAMPLING ORE. 
Ba | Size of Safe Weight in Pounds when Largest Particles are of Size Given 
—£&) Particles. in Second Column. 
| reer ce ee ee aE eae 
eai\S| dea Grade of Richest Mineral Divided by Average Grade. 
Sa) oO) 88 
pa) ae | an 10 50 200 600 1500 2500 
0043} 003 010 025 043 
We ‘0055 .0003 0018 .007 021 053 089 
50| .0100 0017 0095 039 116 291 485 
5.0| 14| .0364 0585 319 1.29 3.90 9.76 16.3 
Fecal F450 2.96 16.13 65.5 195. 494. 823. 
3| 338 | 30.03 | 163.5 663.9 1998. 5000. Soon 
af 75.9 413. 1679. 5054. 12648. 21095. 
1.0 486. 2646. 10746. 32346. 80946. | 140346. 
120] .0043 005 015 038 064 
100) .0055 0005 0027 O11 032 080 134 
50! .0100 0026 0143 058 174 437 727 
7.5| 14| .0364 0878 479 1.94 5.85 14.64 24.5. 
Seeds 1450| 4.44 24.2 98.3 293. 740. 1234. 
9) .338 | 45.0 245. 996. 2997. 7500. 12505. 
4 114. 620. 2519. 7581. 18972. 31643. 
1.0 729 3969. 16119. 48519. |121419. | 210519. 
120| .0043 0005 0027 011 .032 081 135 
100! .0055 0010 0055 022 .068 170 283 
10.5| 50) .0100 0041 0222 .090 272 679 1.133 
4| .0364 1476 804 3.26 9.83 24.59 41.00 
abe 450). °° 7.78 42.35 172.0 517.7 1295. | 2160. 
2) 338 | 78.8 429, 1742. 5245. 13126. 21883. 
5 230. 1250. 5077. 15283. 38247. | 63762. 
<a 1500 3000 6000 15000 30000 | 60000 
~~ 1150} .0036 0798 159 319, 798 Teeth 3.19 
120) .0043 1359 272 | 544) 1.36 2:72 5.40 
100) .0055 284 569 1.138 2.84 | 5.69 | 11.38 
17.6) 50} .0100) 1.14 2.28 4.56 | 11.4 22.8 45.6 
14| .0364| 41.2 82.5 165. 412. 825, 1650. _ 
4| .1450] 2172. 43.46. 8692. 21716. 43461. 86922. 
2| .338 122004. |44037. 88074. | 220038. 1440370. |880740. _ 


34 


as carefully as indicated by the table throughout the entiré 6pération. 

It should also be noted that except in the case of native metals the 
richest minerals are usually more finely divided by crushing than the 
gangue so we seldom have the extreme case provided for by the for- 
mula. 

One of the most difficult things an assayer may be called upon to do 
is to sample such a mill product as vanner concentrates. In these 
the particles of gangue minerals are two or three times the diameter 
of the average rich mineral and good mixing is impossible. The 
material stratifies on the slightest provocation and thé greatest care 
must be taken if the sampling is to be successful. 


Duplicate Sampling. To check the accuracy of the sampling 
operations, we may resort to the process of duplicate sampling or to 
re-sampling. Duplicate sampling in the laboratory should ¢onsist 
in first cutting the entire lot into two portions and then sampling 
each one down separately. The results should check usually within 
one per cent. If not, it indicates either poor mixing and cutting or 
a too rapid reduction of sample. 

Some sampling mills are arranged to allow for taking duplicate 
samples so that they have constant checks on the accuracy of their’ 
sampling operations. The following results of assays made on original 
and re-sampled lots are taken from D. K. Brunton’s paper on Modern 
Practice of Ore-Sampling in the Transactions of the American In- 
stitute of Mining Engineers: and shows how closely such work is 
made to check. 


TABLE VII. RESULTS OF RE-SAMPLING. 


Re-sample Gold. 


Sample Gold. 

Lot No. Ounces per ton Ounces per ton 
3192 6.62 ran 
3198 5.04 5.015 
ete 270 2.67 
3235 3.18 8.16 
3310 1.17 ne 
3324 6.52 ae 
3340 0.71 0.78 
3388 1.70 nee: 
3494 9.24 9.20 

30.52 





3471 30.64 








Finishing the Sample. 
When the sample has been reduced to a few pounds of 40 or 50 
mesh ore, from one to five or more pounds depending on the character 
IT, A. I. M. E. XL. p. 567 (4909) 


a" 7 hue SS Ce 


we a Sr ee 


a ea Pe ee a. ee ee 





39 


of the ore, is ground to 120,150 mesh or finer. This final grinding 
may be done on the bucking board or in any of the numerous forms of 
sample grirniders. As the sample grows smaller more and more care 
has to be taken to prevent contamination or “salting.” A few 
particles of rich ore, which if introduced into the original lot, would 
have had no material effect on the average might easily seriously alter 
the result if allowed to enter the final sample. 

The grinding machines, sieves etc. should be so constructed that 
they may be easily and thoroughly cleaned. Many excellent pulver- 
izers are unsuited for sampling work on account of the labor and 
difficulty of effective cleansing. One of the best methods of cleaning 
the bucking board or sample grinder is to brush out, then grind a 
quantity of some barren material such as sand or crushed fire-brick 
and follow this by a second brushing. All split-shovels, brushes, 
screens and rolling cloths must also be carefully cleaned before use. 

Before commencing the final pulverizing, the sample should be 
thoroughly dried by heating to 100° or 110°C. No greater degree 
of heat than this should be used as there is danger of roasting the 
sulphides or otherwise altering the composition of the ore. 


Size of Assay Pulp. For assay purposes, all ore should be reduced 
to at least 100 mesh and rich spotty ores should be pulverized to 
120 or 140 mesh or finer to insure a fair sample being obtained for the 
final crucible or scorification assay. For a crucible assay using 1 


~ assay-ton an ore may be left coarser than for a scorification assay 


where only 0.1 assay-ton charge is used. If the assayer has difficulty 
in obtaining results checking within one-half of one per cent he may 
well look for the difficulty in the size of the assay pulp. Very often 
a regrinding to a finer size will overcome the difficulty. 

When any portion of ore has been selected as a sample and is to be 
passed through a sieve, it is essential that the whole sample be made 
to pass. The harder portions which resist crushing the longest are 


almost invariably of a different composition from the remainder and 


if rejected render the whole sample worthless. 


Ores Carrying Malleable Minerals. 


In crushing ores containing metallic particles and other malleable 
minerals more or less of those will be left on the sieve as flat scales, 
cylinders or spheres. When such an ore is being sampled the par- 
ticles found on each sieve must be separately preserved and weighed 
and a careful record made of the weight of the original ore and of 
each of the cuts throughout the sampling process. If the pellets 


36 


are gold or silver, they are wrapped in lead foil and cupelled, weighed 
and parted. If of copper, as in the case of an ore containing native 
copper, the weight of the metallic contents is otherwise established, 
perhaps by cleaning in hydrochloric acid and direct weighing or by 
making a fusion as in the copper assay. 

The following example shows how with the above data and the assay 
of the fine ore the assay value of the original sample is calculated. 
It may be either higher or lower than the fine ore. One assumption 
has to be made in all such cases and that is that the reject contains 
the same proportion of pellets as the sample. It need hardly be men- 
tioned that if the proper ratio between size of sample and maximum 
grain has been maintained the above assumption will be born out in 
practice. 


CALCULATION OF ASSAY WHEN ORE CONTAINS COARSE PARTICLES ~ 
OF NATIVE GOLD. 


Data. 


A sample of 23.95 kilo- |- On sieve 25 grams. |- On sieve three grams. 
grams or 23,750 grams was | This yielded 6.2750 grams we ielded 1.6720 grams 
crushed to pass a 40 mesh-| of gold. pa 
sieve. — Through sieve 23,600 |- o Peas sieve 5802 

grams (Loss 125 grams). grams (Loss 20 grams). 
A sample from this of The fine ore assays 1.20 
5825 grams was crushed-| oz. gold per ton. 
| to pass a 120 mesh seive. 


Calculations. Wt. Gold 
Total pellets from 23,750 grams of ore on 40 mesh 6.27500 grms. 
Total 40 mesh ore assuming loss to be same as the 
rest i.e. sample now 23,725 grams. 
Total pellets from 23725 grams on 120 mesh 
23725 











= —— X 1.6720 = 6.81006 
5825 
Assuming all of ore to be crushed through 120 mesh 

and no loss there would be 23,725 ~ — 
= 23,713 grams fine ore (assaying 1.21 oz.) 

Total gold if this = 0.98146 

29.166 rs 

Total gold in original lot | 14.06652 


29.166 : XxX = 23750 : 14.06652 
x = .01727 = Au from 1 assay-ton 
Ore assays 17.27 oz. per ton. 


ne 





CHAPTER IV. 


BALANCES AND WEIGHTS. 


The reliability of every assay or other quantitative determination 
is directly dependent upon the accuracy of the weighing, both of the - 
ore charge and more especially of the resultant product, for example, 
the silver button or the parted gold. Any error made in the weighing 
will, of course, invalidate all the rest of the work regardless of any 
amount of caré which may have been given it. The operator should, 
therefore, familiarize himself with the construction, sensibility and 
operation of his balance before he attempts to do any accurate assay- 
ing. 

A good assay balance with careful and intelligent use is capable 
of weighing to 0.01 milligram or 0.00001 gram. For the most delicate 
assay balances an accuracy of 0.000002 gram is claimed. The neces- 
sity of weighing to this degree of accuracy may be understood when 
it is considered that taking the usual charge of ore, 1 assay-ton, 
(29.166 gms.) or about an ounce, that in weighing the resultant gold 
to the nearest 0.01 milligram the value of the ore is only determined 
to within 20 cents per ton. This is usually sufficiently close, but any 
less degree of accuracy would not be so considered. 

At least three grades of balances are necessary for the fire assay 
laboratory, these are known as flux, pulp, and button or assay bal- 
ances. In large assay laboratories, there are also usually found bullion 
‘and chemical balances as well as separate assay balances for gold and 
for silver. 


Flux Balance. The flux balance for the weighing of fluxes, 
reagents, etc. should be an even balance scale, provided with a remov- 
able scoop-shaped pan capable of weighing 2 kilograms and sensitive 
to 0.2 gram, 


Pulp Balance. The pulp balance for weighing the ore or pulp 
for assay and the buttons from lead assays etc. should be an even 
balance scale. The pans should be made removable and should each 
have a capacity of at least 2 ounces of sand. It should be enclosed 
in a glass case and should be sensitive to one milligram. Such bal- 
ances are usually listed in the manufacturers’ catalogue as prescrip- 


38 


tion balances and may be purchased for $20 or $25. If more than one 
pulp balance is to be obtained, it is well to get one or more having a 
pan capacity of 4 or 5 ounces of sand. For 3 and 1 assay-ton charges 
the two ounce pan is to be preferred as it is easier to transfer ore from 
it to the crucible than with a larger pan. 


Button Balance. The button or assay balance is the most sensi- 
tive balance made. It should be capable of weighing to at least 
0.01 milligram, should be rapid in action, making a complete oscilla- 
tion in from 10 to 15 seconds, and should have stability of poise, that 
is to say that it should be so made that its adjustments will not 
change sensibly from day to day owing to slight changes of temperature 
and atmospheric conditions. The capacity of the assay balance need 
not be large, 0.5 gram maximum is sufficient, but the beam should 
be rigid at this load. ; 

Such a delicate piece of apparatus must be handled with great care 
if good service is expected of it. It should be as far as possible from 
any laboratory or part of the plant where corrosive fumes are being 
evolved, and should be covered when not in use to keep out the dust. 

The balance beam should be as light as possible consistent with the 
necessary rigidity. For this reason the truss frame construction is 
usually adopted, giving the maximum strength with the minimum 
weight. The construction should be such that the two balance 
arms are of equal weight and length and the three knife-edges should 
all lie in the same plane. The material of the beam should be non- 
magnetic for obvious reasons and should have a small coefficient 
of expansion. The knife-edges and bearings should be of agate, 
ground and polished. The knife-edges should be so sharp that a 
strong pocket-lens will show no flatness on the bearing edge. All 
of the metal work of the balance should be protected from attack 
by chemical fumes by some such means as gold-plating or lacquering. 
Lacquering seems to resist chemical fumes rather better than the 
ordinary gold-plating. The construction of the balance should be 
such that the rider may be placed on the zero graduation and used 
from the zero point to the end of the beam. 

The balance must be mounted in such a way that it will be free 
from vibration. Such a support may be obtained by placing the 
shelf on which the balance rests, on one or more posts whieh are set 
in the ground and which come up through the floor without touching it. 


Theory of the Balance. The balance is essentially a light 
trussed beam, supported at its center by a knife edge. At each end 
is hung a scale-pan, both of which should be of equal weight. 





39 





M+m 


M 


Let the three knife-edges A, B, and C be in the same straight line. 
Let AB = BC = 1. Let Gbe the center of gravity of the beam whose 
weight is W. Let the distance of the center of gravity below the 
point of support, BG = 1’. 

With a load of M in each pan there will be equilibrium. Now if 
a small weight (m) be added to the right-hand pan, the balance will 
swing through a small angle 6 and the beam will again come to 
- equilibrium in a new position A’BC’. The condition for equilibrium 
will be obtained by taking the moments of the three forces, M, 
M + m and W about the axis B. This gives the relation: 

MI cos 8 + |’ sin 09 W = (M + m)l cos @ 

m 





The sensitiveness of a balance is usually denoted by the angle 
through which the beam will swing when a small weight usually 1 
milligram, (for assay balance 0.1 milligram) is added to one pan. 
For small angles the tangent and its angle may be taken as equal 
and therefore the expression deduced for tangent 8 above may be 
taken as a measure of the sensitiveness of the balance. 

The equation for tangent shows that the sensitiveness of a balance 
varies: 

(a) Directly as the length of the balance arms. 

(b) Inversely as the weight of the beam. 

(c) Inversely as the distance of the center of gravity below the 

point of support. (Distance BG.) 

The sensitiveness is seen to be independent of the load if the three 
knife-edges are in the same straight line and most balance makers 
‘attempt to approach this condition in making assay balances. When 
B is above AC the sensitiveness is decreased with the load, when B 
is below AC it is increased up to a certain limit, beyond which the 
equilibrium becomes unstable. 

The condition of increased sensitiveness with long beam and small 
weight ((a) and (b) above) conflict, as the longer the beam is made 
the heavier it must be. The length of the arm is also limited by the 


40 


time of swing of the balance which may be considered to be a com- 
pound pendulum. A period of about 10 or 15 seconds is about the 
time required for a complete oscillation. Formerly the long arm 
balances were common but at present the makers are restricting the 
length of the beam to 4 or 5 inches. 

By bringing the center of gravity nearer to the center of support 
the sensibility is increased. As the center of gravity nears the center 
of support, the stability of poise decreases until when the two coincide 
there would be no point of rest and the balance would be unstable 
or ‘cranky.’ The most difficult thing to obtain is a balance with 
great stability and extreme sensibility. It is obtained by making 
the beam as light as possible and then keeping the center of gravity 
sufficiently below the center knife edge to give the necessary stability. 
Most high grade balances are provided with a screw-ball or sliding- 
weight so that the center of gravity may be adjusted. If the balance 
lacks stability, 1.e. is cranky and over-sensitive, both of those con- 
ditions may be reduced by lowering this weight and thus lowering 
the center of gravity of the system. 

In the above discussion the assumption has been that the arms of 
the balance were equal. Modern high grade balances usually ap- 
proach very closely to this condition. The process of double weigh- 
ing serves to eliminate, however, any error in weighing that may be 
due to such inequality. Call the observed weight of the body as 
weighed in pan A, W’, and that in pan C, W”. Then W the true 
weight is found as follows: 

P= /wew" vidas 
when W’ and W” are nearly equal W = V2 (W’+ WwW’) 


General Directions for the Use of the Balance 


1. Note the maximum load the balance will carry and do not exceed 
this. 

2. The balance should be put into action by gently lowering the 
beam onto the knife edges. 

3. Have the amplitude of swing not more than 3 or 4 divisions 
each side of the center. 

4. Arrest the swinging of the balance when the pointer is at’ the 
center of the scale. 

5. Turn the balance out of action before adding to, or taking 
weights from the pan. | 


41 


6. When not in use raise the beam off the knife edges and leave 
the rider on the beam. 

7. The final weighing must be made with the case closed, 

8. Weights should be placed only in the box or on the scale pan and 
should be handled only with the forceps. 

9. Record the weight of the substance; first by noting the weights 
which are absent from the box, second by checking off each weight 
as it is put back in the box. 


Weighing. 


Brush off the balance pans and if necessary clean off the front 
plate of the balance. Adjust if necessary and try the adjustment 
each time you have any weighing to do. In putting the balance 
into action it may start swinging slightly of its own accord. If it 
does not, set it swinging by gently fanning one pan with a motion 
of the hand, or by lifting the rider for an instant and then putting 
it back on the beam. A device such as a medicine dropper may be 
used for starting the balance swinging by blowing with it gently on 
one pan. If the balance is started swinging by fanning with the 
hand it should be allowed to make one or two complete oscillations 
before a reading is taken, to prevent air currents from interfering 
with the normal swing. 

In reading the position of the pointer on the ivory scale, arrange 
always to have the reading eye in the same position relative to the 
ivory scale, that is in a plane perpendicular to the scale and passing 
through the center graduation. A mark may be made on the glass 
door by which to line up the eye before each reading. 

Each silver bead should be placed on its side on a small anvil, 
hammered and then brushed before it is weighed. 

To transfer the gold from the parting cup to the scale-pan take the 
seale-pan with the forceps and place on the front part of the glass 
mounting base. Gradually invert the parting cup over it tapping it 
gently. The gold should all slide into the pan. Any particles 
adhering to the cup may be detached by touching gently with the 
point of the geEceDs or by means of a small feather trimmed to a 
point. 

Before eephite the gold, examine it carefully to see if it is clean 
and remove any foreign matter if present. 

For ordinarily accurate commercial work the weighing of the gold 
and silver is done by the “‘method of equal swings’’ using the rider for 
the final weighing. For extreme accuracy as for instance in the cal- 


42 


libration of weights, the weighing is done by ‘‘deflection”’ also called 
the ‘‘method of swings.” 


Weighing by Equal Swings. First of all the balance is adjusted 
by the star wheel or preferably by the adjusting rider, if one is pro- 
vided, until the needle swings exactly the same on each side of the 
center, reading always in the same order, say from left to right. For 
accurate gold weighing it will be necessary to estimate tenths of divis- 
ions on the ivory scale. 

Put the -substance to be weighed on the left-hand pan and add 
weights to the right-hand pan until within a fraction of a milligram 
of the weight of the substance. The application of the weights should 
be done in a systematic manner, starting with one which is estimated 
to be too large. If too large, it is removed and the next smaller weight 
tried and so on, always working from larger to smaller weights un- 
til within 1 milligram of the true weight. 

In trying any weight have the balance off the knife edges, put the 
weight in the pan and gently turn the balance key until the pointer 
inclines slightly to one of the other side. This swing of only one or 
two divisions should indicate immediately whether the weight on 
the pan is too much or too little. Again turn the balance out of 
action before making any change of weight. 

When within a fraction of a milligram of the correct weight, the 
right hand or weighing rider is shifted about until on putting the 
balance into action the needle does not move very decidedly in one 
or the other direction. Then set the beam swinging two or three 
divisions each side of the center. If it does not swing exactly evenly 
arrest the swing and change the position of the rider and try again. 
Repeat until the needle swings exactly the same as when adjusted. 
After becoming familiar with the balance two or three trials only of 
the rider will be necessary. 

The weight of the substance is found from the sum of the weights 
on the pan plus the fractional parts of a milligram indicated by the 
position of the rider on the beam. 


' Weighing by Method of Swings. First, determine the position 
of rest of the pointer under zero load by noting the position of the 
pointer at the extreme swing on each side, taking 3, 5 of a greater 
odd number of consecutive readings. Call the center division zero 
and count divisions and estimate tenths to each side, calling those 
to the left of the center — , and to the mnght +. 





43 


Example. 











Left Right 
~3.9 3.6 
~3.7 3.4 
—3.90= 2 [7.0 
3 |-11.1 Pa 
~3.7 
+ 3.5 
Do 2, 
— 1 Point of rest. 


Or the point of rest would be 0.1 division to the left of the center, 
Call the point of rest under zero load r. Place the object to be 
weighed on the left-hand pan and weights on the right-hand pan 
until equilibrium is nearly established. With the rider determine 
the weight to the next smaller 0.1 milligram. Set the beam swinging 
as before and find the position of rest for the pointer. Call it r.’ 
Shift the rider to the right one whole division ( = 0.1 mg.) so as to 
bring the point of rest on the opposite side of r, and find the position 
of rest again, call it r’’. The tenths of a milligram to be added to 
the weights and rider reading when r’ was found is then 
r’—r 
ie 
For instance let the weights and rider reading be 27.4 mg. and 
let r’ = — 14 andr’ = + 1.6 
ae ate | Ca 
then s = e] 1.4 +4 0.1 £ als ~ 40,43 
r=-r —~14.— 1.6 = 3.0 
and the true weight would be 27.4 + 0.04 = 27.44 még. 

Another method of weighing by “deflection,” requiring a knowledge 
of the sensibility of the balance is as follows. Suppose that a weight 
of 0.10 milligram will cause a deflection of the point of rest 2.0 divi- 
sions on the ivory scale. Adjust the balance so that the point of rest 
with no load corresponds to the zero of the ivory scale. Place the 
substance to be weighed in the left-hand pan and again determine 
the point of rest. Suppose this is 1.2 (divisions). Then the weight 
of the substance is 0.06 milligrams. With a good balance this is a 
rapid and accurate method for small amounts of gold but is not very 
- commonly used. 


Weighing by No Deflection. A third method of weighing 
called weighing by ‘‘no deflection” is sometimes employed for rough 


44 


work. It consists in applying the necessary weights and then shift- 
ing the rider until the needle shows no deflection when the balance 
is lowered gently onto the knife edges. This method disregards 
friction and inertia and is not as accurate as the two previously des- 
cribed methods. 


Weighing by Substitution. This method of weighing is the one 
usually adopted for the standardization or adjustment of weights, 
as it avoids any possibility of error due to inequality of arms. It 
consists simply in placing the substance to be weighed on one pan, 
counterbalancing it with weights placed on the other pan, and then 
removing the substance and adding standard weights until the bal- 
ance is again in equilibrium. The weight of the substance being 
obtained from the substituted weights. 


Check Weighing. Students are advised to check all gold weigh- 
ings in the following manner. Weigh and record weight of each 
duplicate, then place both buttons or prills on one seale-pan and ob- 
tain the weight of the two. Compare this with the sum of the weights 
_ obtained in the separate weighings. The figures should check within 
0.01 of 0.02 milligram. If not, either one of the weighings is at fault 
or some of the weights are in error. Check weighing also serves 
to double the accuracy of the assay and is practised by many of the 
best assayers. 


Adjusting and Testing an Assay Balance. 


Levelling. Level the balance by adjusting the footserews and 
by observing the plumb-bob of level. Be sure that it rests firmly 
on the table or other support so that it will not be moved during the 
test. See that the beam, scale-pans and hangers are in their proper 
places and not out of normal position due to previous careless usage. 


Equilibrium. Lower the beam carefully until the agate knife- 
edges rest on the agate supports. This motion and the reverse one 
must be gentle to prevent injury to the knife-edges and also as any 
shock or jar will tend to change the adjustments. Adjust the bal- 
ance so that it swings equally on each side of the center. A nut 
on one or each end of the beam, a star-wheel, a small projecting 
piece of metal or ‘‘flag’’ revolving on a vertical axis at the middle of 
the beam, or preferably an extra rider, constitutes the attachment for 
this adjustment. If this adjustment cannot be made but the balance 
on starting to one side or the other continues to swing in that direc- 
tion with increasing velocity, it is in unstable equilibrium, and the 





45 


center of gravity must be lowered until the proper equilibrium is 
obtained. 


Time of Oscillation. Set the balance in motion and note the 
time of one complete oscillation, i.e. swing from one extreme to the 
other and back again. For the modern 4 or 5 inch beam assay bal- 
ance this should be from 10 to 15 seconds. If much faster than this 
the balanee will probably not be very sensitive. If much slower 
than this each weighing will take a correspondingly longer time and 
the balance may lack stability. 


Stability. After each of the tests the beam should be lowered 
and the adjustment of the balance noted. If it no longer swings 
equally on each side of the center, due care having been taken to 
avoid disturbing any of the adjustments, it lacks stability. This may 
be due to its adjustment to a too great degree of sensitiveness which 
can be overcome by lowering the center of gravity of the system by 
means of the screw ball, or it may be due to defect im construction, 
arms of unequal length, etc. in which case it can not be remedied. 


Resistance. If the knife-edges are dull or the supporting surfaces 
rough the frictional resistance to swinging will be considerably and 
the diminution in amplitude of swing will be rapid. Note the posi- 
tion of the pointer on the scale at the extremes of several successive 
swings. The difference between successive readings on the same side 
will show the diminution in amplitude due to friction and to resistance 
of the air. This should not exceed 0.2 to 0.4 of a division in a good 
assay balance. The horizontal section of the beam and the area 
of the pans and other projecting parts should be as small as possible 
to reduce the air resistance. 

Let the balance swing until it comes to rest and read the position 

of the pointer, lift the balance off of the knife-edges and repeat sev- 
eral times. The positions should not differ by more than 0.05 of a 
division. This indicates a flatness of the knife-edges or a roughness 
of the supporting surfaces. If the beam is exceedingly slow in com- 
ing to rest this test is unnecessary. 


Sensibility. The sensitiveness is defined by physicists as the 
angle moved through by the beam when 1 milligram excess weight 
is added to one pan. The angle being proportional to the number of 
scale divisions passed over, this latter is taken as a measure of the 
sensitiveness. : 

From a practical point of view the sensitiveness is the smallest 
difference in weight which the balance will indicate. Thus, when we 


46 


say that a balance is sensitive to 0.01 milligram we mean that 0.01 
milligram added to one pan will cause a noticeable difference in the 
swing or in the zero! point. With the usual width of graduation of 
the ivory seale 0.05 inches, the pointer should swing over at least 
one-fifth of a division for each 0.01 milligram added to the scale-pan, 
if the balance is to be termed sensitive to 0.01 milligram. 

The writer has never seen described a satisfactory practical method 
of comparing the sensibility of different assay balances, but has used 
the following method based on the commonly accepted space of 0.05 
inches, between the graduations on the ivory scale. 


Procedure. Determine the zero point with any given load on 
the pans, then add 0.10 milligram to the weight on the right-hand 
side of the beam using the rider and again determine the zero point. 
Multiply the difference between these two positions (expressed in 
terms of units of graduation) by the distance apart of the graduations 
in hundredths of inches and divide by 5. 

Comparative sensibility = (deflection of zero due to 0.1 mg.) = 
(space between graduations in hundredths of inches) 


5 


Example. Deflection 1.2 divisions, space between graduation 
0.05 inches. 








1.2 Xx 9 


Comparative sensibility = 1.2 which means that this 


balance will weigh to something better than 0.01 milligram. 


To Test Equality of Arms. Adjust balance to swing evenly 
with no load and then place equal weights on each pan equivalent 
to the full load of the balance. If the pointer does not now swing 
evenly the arms are of unequal length. 


To Determine if Knife Edges are all in Same Horizontal 
Plane. Adjust balance and determine sensibility with no load. 
Then place full load on each pan and again determine sensibility. 
When the three knife edges are in the same place there should be no 
change of sensibility with anything up to the full load of the balance. 
When the full load of the balance is not known the sensibility should 
be determined for gradually increasing loads and a curve of sensibility 
drawn. If the three knife edges are in the same plane this curveshould 
be a straight line up to the point where the beam begins to be deflected 
by an overload. 


1 zero point = position of the point of rest. 


‘ 47 


Weights. 


For the three balances above described we require four sets of 
weights, as follows:— 

For the flux balance we should have a block containing weights 
from one kilogram to one gram. These weights need not be extremely 
accurate. 

For the pulp balance two sets are necessary, gram and assay-ton 
_ weights: gram weights, from 20 grams to 10 milligrams for weighing 
| argols and ore for lead, copper and tin assays as well as the buttons 
_ from the same: assay-ton weights, 2 A. T. to zs A. T. for weighing 
ore, matte, speiss and lead bullion for the gold and silver assay. 

For the button balance is required a set of milligram weights of the 
utmost accuracy, from 1 milligram up to 500 or 1000 milligrams. 
These are preferably made of platinum as an absolutely non-corrosive 
weight is imperative. Rzders are used for determining fractions 
of 1 milligram. Riders are made of fine platinum or aluminum 
wire and are usually made to weigh 0.5 milligram or 1.0 milligram. 
The balance beam is divided usually into 100 spaces. each side of the 
center and then when a 1 mg. rider is used each space represents 0.01 
milligram. 

For many balances a rider with a diamond shaped Joan known as 
the Thompson rider is to be preferred. Its principal advantaye is 
due to its property of always hanging in a vertical position when on 
the rider arm. Even if it falls over to one side when on the beam it. 
will slip back to the vertical position when lifted by the rider arm. 
The diamond-shaped loop prevents it from swinging or twisting 
- around on the rider carrier and permits the rider to be placed squarely 
‘on the beam. 





























Multiple Rider Attachment. Some of the balance makers are 
now supplying on demand what is called a multiple rider attachment 
designed to do away with the use of the smaller weights. It consists 
of a carrier supplied with a number of riders of different weights for 
‘instance 1, 2, 3, 5, 10, 20, 30 milligrams, so arranged that any of all 
may be placed on a support provided for the purpose, and which is 
equivalent to placing flat weights of the same value in the pan. 

The advantages claimed for this device are a saving in the wear 
and tear of weights, as the small flat weights frequently handled by 
forceps become broken and inaccurate whereas there is practically 
‘no wear on the riders so that they will maintain their original weight 
almost indefinitely. A second advantage claimed is a saving in time 
as with this attachment the weights may be manipulated much 


48 


quicker than can the flat weights with forceps. It is not necessary 
to open the door of the balance in weighing a button under 40 or 50 
milligrams and this alone is a saving of some time and also allows 
all air currents to subside before the final reading is made. 


Assay-Ton Weights. The assay-ton system of weights was 
devised by Professor C. T. Chandler of Columbia University, to 
facilitate the calculation of results of gold and silver assays. In 
the United States and Canada the results of ore assays are reported 
in troy ounces per ton of 2000 pounds avoirdupois. With the ordin- 
ary system of weights a tedious calculation would have to be made 
for each assay with the possible mathematical errors. 

The basis of the assay-ton system is the number of troy ounces 
(29.166 + ) in one ton of 2000 pounds avoidrupois. The assay-ton 
is made to weigh 29.166 grams. Then 

l'ton av : 1 oz. troy. 2:1 AcT> |) milion 
Therefore using one assay-ton of ore the weight of the silver or gold 
in milligrams gives immediately the assay in ounces pér ton. 


Calibration of Weights. The weights supplied by the makers: 
cannot always be relied upon and even originally perfect ones are 
subject to changes of weight due to wear or accumulation of dirt. 
Therefore it behooves the assayer to occasionally check his weights 
and to determine the correction to be applied to the marked value. 
This requires the use of a standardized weight which should be care- 
fully preserved and used for this purpose only. 

The method of swings should be used and the weighing is fone by 
deflection after the sensibility (deflection of the zero for 0.1 mg.) 
has been determined. First determine the position of rest, and the 
sensibility with no load, with 100, 250 and 500 milligram loads re- 
spectively. The sensibility should not vary much throughout this 
range. The method to be followed can be understood from the 
following examples. 


Calibration of a Set of Assay Weights. 


Designate each weight by its marked value in the parenthesis and 
when there are several of the same value note some peculiarity by 
which they. may be designated. The weights in the set, marked in 
milligrams are: 

(500) = a, (200) = b, (100) = ¢, (1009 =:d) (50) = eee 
(10) = g, (10) = h, (10) = 6) +2) +@)4+@Q) =i 
The weight (100) is compared with the standard 100 milligram 


49 










weight and the weights are then compared among themselves by the 
method of swings. The letters represent the true values. 

In calibrating the weights from 100 milligrams to 10 milligrams, 
observations should be made on the following combinations: 


Left Hand Pan Right Hand Pan 
(100) 100 mg. standard 
(100) pre (20) 410) (10) 4--(5)-- (2), + (2) 
ul) 
(50) Pee 0) 010!) 5) 2) 29 (1) 
(20) GLO) (10") 
The recorded observations are as follows:— 
100 mg. = c¢ — 0.020 mg. 

¢ Peete eth +i 0/190 

e afte t+h+i + 0.020 

f =g+th + 0.040 

g sah + 0.015 

i + 0.040 

Solving these equations | 

I=} 
Dei + 0.040 
eer + 0.055 
iva + 0.135 
@ = 5i + 0.450 
c= 101 + 0.870 
ec = 100.020 mg. 


_ From these last two values of ¢ 
101 = 99.150 mg. 
P= 9.915 me. 
Substituting this value for i in the above equations we find the 
a following values for the other weights of the set: 


Designation Actual Weight Correction ‘ to marked value 
c = 100.020 + .020 mg. 
e= 50.025 020s 
f= 19.965 — 035 “ 
g= 9.970 — .030 « 
h= . 9.955 — 045 “ 
ae 9.915 — 085 “ 


The smaller weights in 7, may be calibrated in a similar manner. 
_ The large weights (100), (200) and (500) may be standardized by a 
simple modification of the above. 

The process is made much simpler by Baers a complete set of 


1 A + correction means that the weight is heavier than the normal value. 


50 


standard weights which are very carefully handled and kept solely 


for standardizing purposes, and these the larger assay offices usually 
have. 


Testing Riders. Every new rider should be tested before use as 
they often vary 0.01 or 0.02 milligrams from their supposed value. 
If too heavy, a little bit at a time may be cut off with a pair of scissors 
until they come down to the standard. 


CHAPTER V. 


CUPELLATION. 


In every assay of an ore for gold and silver we endeavor to use such 
fluxes and to have such conditions as will give us as a resultant two 
products :— 

Ist. An alloy of lead with practically all of the gold and silver of 
the ore and as small amounts of other elements as possible. 

2nd. A readily fusible slag containing the balance of the ore and 
fluxes. | 

The lead button is separated from the slag and then treated by a 
process called cupellation to separate the gold and silver from the 
lead. This consists of an oxidizing fusion in a porous vessel called 
a cupel. If the proper temperature is maintained the lead oxidizes 
rapidly to PbO which is partly (98.5 * per cent) absorbed by the cupel 
and partly (1.5 per cent) volatilized. When carried to completion 
the gold and silver is left in the cupel in the form of a by+* = 

The cupel is a shallow, porous dish made of bone-ash, Portland 
cement, magnesia or other refractory and non-corrosive material of 
spongy texture. The early assayers used cupels of wood ashes from 
which the soluble constituents had been leached. Agricola writing 
in about the year 1550 mentions ashes from burned bones. Ashes 
from deers’ horns alone he pronounces best of all; but the use of these 
antedate his time and he states that assayers of his day generally 
make the cupels from the ashes of beech wood. 

To-day it is thought that the bones of sheep are the best for cupels. 
These should be cleaned before burning and as little silica as possible 
introduced with the bones. It is important not to burn the bones 
at too high a temperature as this makes the ash harderand less absorb- 
ent. It is also advisable to boil the bones in water before burning 
them as this dissolves a great part of the organic matter which if 
burned with the bones yields sulphates and carbonates of the alkalies. 

Properly burned sheep bones will yield an ash containing about 
90% calcium phosphate, 5.65% calcium oxide, 1.0% magnesium 


oxide, and 3.1% calcium fluoride. Ordinary commercial bone-ash 


also contains more or less silica and unoxidized carbon. If more 

than a fraction of a per cent of silica is found in bone-ash, it is evidence 

that sufficient care has not been taken in cleaning the bones, and cupels 
1 Liddell, E. & M. J. 89, pp. 1264. June 1910. 





52 


made from such bone-ash aremore likely to crack during cupellation, 
resulting often in the loss of small buttons. If the bone-ash shows 
black specks it is an indication of insufficient oxidation and the assayer 
should allow the cupels to stand for some time in the hot muffle with 
the door open before using. Carbon is an undesirable constituent 
of cupels as it reacts with the lead oxide formed giving off CO and 
CO: which may cause a loss of the molten alloy due to spitting. 
Bone-ash for cupels should be finely ground to pass at least a 40- 
mesh screen and the pulverized material should consist of such a 
natural mixture of sizes as will give a solid cupel with enough fine 
material to fill interstices between coarser particles. Opinions differ 
as to the best size of crushing for bone-ash and this will depend no 
doubt upon the character of the material. Bone-ash, the screen 
analysis of which follows, has, however, yielded particularly good 
cupels. 
TABLE VIII. SIZE Oe SORE ASH. 











oe 











Size Mesh Size mm. ne Cent. L Wee 
On 40 0.380 9 Pou Cents: 
Through 40“ 60 0.244 is as 
60 “ 100 0.145 gees ee 
+ 100 “ 156 0.098 10: Piva 
a 150 50 4c SS 











With cupels made from this bone-ash it was possible to obtain 
losses not exceeding 1.60 per cent using 100 mgs. of silver and 25 
grams of lead, while with some other lots of bone-ash containing smaller 
proportions of 150 mesh material it was found impossible to keep the 
losses below 2.0 per cent. 


Making Cupels. Cupels are made by moistening the bone-ash 
with from 8 to 20 per cent of water and compressing ina mold. The 
bone-ash and water should be thoroughly mixed by kneading, and 
finally it should be sifted through a 10 or 12 mesh sieve to break up 
the lumps. Some authorities recommend adding a little potassium 
carbonate, molasses or flour to the mixture, but with good bone-ash 
nothing but pure water need be added. The mixture should be just 
sufficiently moist to cohere when strongly squeezed in the hands, but 
not so wet as to adhere to the fingers or to the cupel mold. Twelve 


per cent of water by weight is about right; but the amount used will — 


depend somewhat on the bone-ash and on the pressure used in forming 
the cupels. The greater the pressure the smaller the amount of water 
which may be used. It is better to err on the side of making the mix- 
ture a little too dry than too wet. 





53 


The cupels may be molded either by hand or machine. The hand 
outfit consists of a ring and die. The ring is placed on the anvil 
and filled with the moist bone-ash, the die is inserted and pressed down 
firmly. It is then struck one or more blows with a heavy hammer or 
mallet, turning the die after each blow: finally the cupel is ejected. The 
cupels are placed on a board and dried slowly in a warm place. The 
amount of compression is a matter of experience and no exact rule 
for it can be given; but it may be approached by making the cupels 
so hard that when removed from the mold they are scratched only 
with difficulty by the finger nail. One man can make about 100 
-cupels an hour using the hand mold and die. 

Several types of cupel machines are on the market. One machine 
has a compound lever arrangement which gives a pressure on the 
cupel equal to twenty times that applied to the hand lever, and by 
adjusting, different degrees of compression may be obtained. These 
machines have interchangeable dies and rings so that different sizes 
of cupels may be made. The rated capacity of this machine is 200 
cupels an hour. Another machine made by the same company has an 
automatic charging arrangement. This machine is claimed to have a 
capacity of 600 cupels an hour. Cupels should be uniform in hardness 
and it would seem that with a properly designed machine a more 
uniform pressure could be obtained than by the use of hammer and 
die. Some assayers however, still prefer hand-made cupels. 

Cupels should be air dried for several days at least before use. Most 
assayers make them up several months in advance so as to insure 
complete drying. They should not be kept where fumes from part- 
ing can be absorbed by them as the CaO present will be converted into 
Ca(NO3)2. This compound is decomposed at the temperature of 
cupellation and may cause spitting of the lead button. 

Cupels should not crack when heated in the muffle and should be 
sufficiently strong so that they will not break when handled with the 
tongs. Good cupels give a slight metallic ring when struck together 
after air-drying. It is best to heat cupels slowly in the muffle as this 
lessens the chance of their cracking. 

A good cupel should be perfectly smooth on the inside and of the 
right porosity. If it is too dense, the time of cupellation is prolonged 
and the temperature of cupellation has to be higher, thus increasing 
the loss of silver. If it is too porous it is said that there is again danger 
of a greater loss due to the ease with which small particles of alloy 
can pass into the cupel. The bowl of the cupel should be made to 
hold a weight of lead equal to the weight of the cupel. 

The shape of the cupel seems to influence the loss of precious 


54 


metals. <A flat, shallow one exposes a greater surface to oxidation 
and allows of faster cupellation, it also gives a greater surface of con- 
tact between alloy and cupel, and as far as losses are due to direct 
absorption of alloy, it will of course increase these. The writer using 
the same bone-ash and cupel machine, and changing only the shape 
of the cupel has found shallow cupels to give a much higher loss of 
silver. In doing this work it was found harder to obtain erystals 
of litharge with the shallow cupel without freezing, and it was very 
evident that a higher cupellation temperature was required for the 
shallow cupel. The reason for this is that in the case of the shallow 
cupel the molten alloy is more directly exposed to the current of air 
passing through the muffle and consequently a higher muffle tempera- 
ture has to be maintained to prevent freezing. T. K. Rose’ also 
prefers deep cupels on account of smaller losses. French found shal- 
low cupels less satisfactory on account of sprouting. 


Cupellation. The muffle is heated to a bright red and the cupels, 
weighing about one-third more than the buttons which are to go in 
them, are carefully introduced and allowed to remain for at least 10 
minutes in order to expel all moisture and organic matter. During 
this preliminary heating the door to the muffle is ordinarily kept 
closed, but if the cupels contain organic matter it is left open at first 
and then closed for five minutes or so before the buttons are intro- 
duced. 

When all is ready the buttons are placed carefully in the cupels 
and the muffle door again closed. If the temperature of the muffle 
is correct and the cupels are thoroughly heated, the lead will melt 
at once (326° C.) and become covered with a dark gray or black scum. 
This should di-appear in the course of a minute or two when the but- 
tons become bright and are said to have ‘‘opened up” or “uncovered.” 
If the buttons are practically pure lead this scum disappears when 
the alloy reaches a temperature of about 850° C. When the buttons 
have uncovered, the door of the muffle is opened to admit a plentiful 
supply of air to promote oxidation of the lead, while at the same time 
the temperature is reduced until feather-like crystals of litharge begin 
to form on the cupel just above the lead. If copper, nickel, cobalt, 
iron, etc. are present, the temperature of uncovering and also that 
required for cupellation will be higher, and it may be impossible to 
obtain litharge crystals. When air is admitted to the muffle after 
the ‘‘uncovering”’ the lead becomes lustrous and emits fumes. The 
lustrous appearance is caused by the flame of burning lead. The 

1K. & M. J. 80 pp. 934. Nov. 1905. 





55 


vapor is that of lead oxide. After cupelling has proceeded for a few 
minutes, a ring may be seen around the cupel just above the surface 
of the metal which is caused by the absorbed litharge. If the tempera- 
ture is right for cupelling this will appear to be only very dull red, 
if bright red, the temperature is too high. The color of the alloy 
itself will be much brighter than that of the absorbed litharge, as it 
is in fact much hotter than the cupel or surrounding air, due to 
the heat generated by the rapid oxidation of the lead. Next to the 
formation of abundant litharge crystals, the appearance of the ab- 
sorbed litharge is the best indication of proper cupellation temperature. 

The minimum temperature at which cupellation will proceed has 
been somewhat of a disputed point owing largely to a difference in 
conception of the process and involved conditions. At least three 
methods of measuring the temperature have been advanced, i.e. 
one experimenter held his pyrometer junction one-quarter inch above 
the alloy in the cupel, another placed the junction inside the cupel 
while a third measured the temperature of the alloy itself. Accord- 
ing to Fulton ' the alloy itself must be between 800 and 850° C. 
Litharge melts at 884° (Mostowitch), 906° (Bradford), but passes 
through a. pasty stage before becoming liquid. It would seem that 
the cupel itself must be maintained above the melting point of lith- 
arge in order to allow of absorption. At any event the cupel is much 
hotter than the space around it in the muffle due partly to the heat 
generated by the oxidation of the lead and partly because resting as 
it does on the floor of the muffle its interior portion becomes heated 
by conduction through the muffle floor on which it stands. Brad- 
ford * found 906° C. as the minimum cupel temperature which would 
permit of absorption of litharge. Lodge® found for silver cupel- 
lation with a moderate draft the muffle temperature (taken 4 inches 
above the cupels) should be between 650° and 700° C. 

If the temperature is exactly right feather-like crystals of litharge 
form on the sides of the cupel above the lead. In cupelling for silver 
the temperature should be maintained so that these crystals are ob- 
tained on at least the front half of the cupel, and as the button grows 
smaller they should follow it down the side of the cupel leaving how- 
ever a slight clear space around the button. If the temperature is 
getting too low for the cupel to absorb the litharge, the crystals begin 
to form all around and close to the lead in the cupel, and soon a pool 
of molten litharge is seen forming all around the annular space be- 

1 Western Chemist and Metallurgist, IV. pp. 31. Feb. 1908. 


2 Jl. Ind. & Eng. Chem. 1 pp. 181. 
3 Notes on Assaying, pp. 62. 


56 


tween the lead and the cupel. If the temperature of the cupel is — 
not quickly raised this pool increases in size and soon entirely covers 
the lead and then solidifies. When this occurs the button is said to 
have ‘‘frozen’’, although the lead itself may be liquid underneath. 
Frozen assays should be rejected as the results obtained from them, 
by again biinging to a driving temperature, are usually low. If. 
the freezing is noticed at the start it may be arrested by quickly 
raising the temperature of the cupel in some way, i.e., by closing 
the door to the muffle, opening the draft, putting a hot piece of coke 
in front of the cupel, ete. 

Beginners have difficulty in noting the first symptoms of freezing, 
but all should be able to see the pool of litharge starting. This gives 
the appearance and effect of oil, for if the cupel is moved the button 
slides around as if it were greased. . 

Toward the end of the cupellation process the temperature must 
be again raised, because the alloy becomes more difficultly fusible 
as the proportion of silver in it increases, and in order to drive off 
the last of the lead a temperature of about 900° C. should be reached. 
The temperature should not be raised so high as to melt the erystals 
of litharge, for if this is done too great a loss of silver results. 

As the last of the lead goes off, the button is covered with a brilliant 
film (play of colors) and appears to revolve. The colors disappear 
shortly, the button becomes dull and after a few seconds appears 
bright and silvery. This last phenomena is called the ‘“‘brightening ” 
As soon as this occurs the cupel should be immediately drawn from — 
the muffle far enough to have the button solidify. As it solidifies 
it will ‘“‘flash’”’ or “‘blick,” i.e., suddenly emit a flash of light due to 
the release of the latent heat of fusion, which raises the temperature 
very much for a short time. 

Cupels containing large silver buttons should be drawn to the front 
of the muffle until they chill, and just as the button is about to solidify 
a very hot cupel is placed over them and allowed to stand for several — 


minutes, after which they are slowly withdrawn from the muffle. — 


If this precaution is not taken, the buttons may “sprout” or “spit.” 
This is caused by the sudden escape of oxygen which is dissolved in 
the molten silver and expelled when the button solidifies. If the 
button is allowed to solidify rapidly, a erust of solid silver forms on 
‘the outside, and as the central part solidifies this erust is violently 
ruptured by the expelled oxygen, giving a cauliflower-like growth on 
the button and causing particles of silver to be thrown off. As a 
consegeunce the results obtained from sprouted buttons are unreliable. 
Buttons containing one-third or more of gold will not sprout even if 


57 


rapidly withdrawn from the muffle. Sprouting is said to be an evi- 
dence of the purity of the silver. 

The silver button should appear smooth and brilliant on the upper 
surface, silver-white in color, and spherical or hemispherical in shape 
according as it is small or large. It should adhere slightly to the 
cupel and appear frosted on the under surface. If the button is smooth 
on the bottom and doés not adhere to the cupel it is an indication 
of too low a finishing temperature and will always contain lead. 
If it has rootlets which extend into cracks of the cupel the results 
are also to be taken as unreliable, as some of the silver may be lost in 
the cupel. : 

Very rich alloys of gold and silver have a peculiar mottled appear- 
ance after cupelling begins. Oily drops of litharge appear and move 


4 about on the surface of the alloy and finally run down the side of the 





convex surface and are absorbed by the cupel. This appearance is 
characteristic and once seen is again easily recognized. It may be 
seen toward the end of cupellation with any alloy containing much 
precious metal and is an indication of the approach of the end and a 
reminder that the temperature should be raised to insure driving off 
the last of the lead. 


First Exercise. Practice in Cupellation. 


Procedure. Take from 0.10 to 0.20 grams of silver, but do not 
waste time in weighing and wrap in 25 to 30 grams of sheet lead. 
Prepare two or three of these and cupel one at a time in order to get 
familiar with the operation, and with the correct temperature. To 
study the end phenomena “play of colors,’ “brightening,” “‘blick”’ 
etc., the same of a larger amount of silver may be used with a smaller 
amount of lead, say 10 grams. 

Have the muffle at a bright red, be sure that the cupels are dry and 
then heat gradually until they are red through. Allow at least ten 
minutes for this. Be sure that the cupels weigh more than the lead, 
and that the bowl is sufficiently large to contain the melted alloy. 
Have a row of extra cupels in front of those which are to be used and 
keep them there throughout the process. Keep the door to the muffle 
closed and when the cupel is red throughout and heated to about 
850° C. place the packet of lead and silver carefully in the cupel and 
close the door to the muffle so that the lead will fuse as quickly as 
possible. As soon as the assay begins to “‘drive,”’ note the time, open 
the door of the muffle and lower the temperature of the cupel by pull- 
ing it forward in the muffle, checking the fire if necessary, or by plac- 


58 


ing cold scorifiers, etc. around the cupel. Continue to reduce the 
temperature until feather crystals of litharge are seen forming on at 
least the front half of the cupel. Then continue the cupellation at 
this temperature. Finally finish the assay at a somewhat higher 
heat, increasing the temperature by moving the cupel back in the 
muffle, by starting up the fire, or by shutting off some of the cold air 
supply by partly closing the door to the muffille. If the cupels are 
running very cold it will be necessary to start raising the temperature 
some minutes before the end. The fire should be under good control 
at all times. As soon as the cupellation is finished remove the assay 
carefully from the muffle to avoid sprouting. AIl assayers agree that 
the best results are obtained by having a hot start, a cold drive, and 
a higher heat again at the finish. 


Notes) 1. When doing a large number of cupellations at one time the buttons 
should be charged in order of their size, 1. e. , largest first, so that all may start driv- 
ing together. 

A skillful assayer with a large muffle can run as many as 50 cupellations at one 
time and obtain feather crystals on all. 


2. When doing a large number of cupellations at one time the cupels are not moved ~ 


about after the lead is put in but the temperature is regulated by means of the draft 

and firing and by the use of coolers, (cold scorifiers, cupels, crucible covers, ete.) 

Hs are put in toward the back of the furnace and replaced as soon as they become 
eated. 

3. Bear in mind that although the temperature of the muffle may be as low as 
650° or 700° C., the cupel itself should be shghtly above the freezing point of litharge 
to allow of its being absorbed. It has been found best therefore to protect the body 
of the cupel itself from the draft through the muffle by placing an extra row of cupels 
or a low piece of fire-brick in front of the first row of cupels. 

4. The cupel should be withdrawn from the muffle far enough to cause the bead 
to solidify as soon as the brightening has taken place otherwise a loss of silver ensues. 

5. Besides gold and silver the button may contain platinum, palladium, rhodium, 
iridium, ruthenium, osmium, and iridosmium. 

6. When gold is present in considerable amounts (33%) the buttons will not sprout 
even if taken directly out of the muffle. 

7. Certain elements, notably platinum and tellurium, give the surface of the 
button a frosted appearance. 

8. When the finishing temperature is too low the buttons solidify without blicking. 
They retain lead and have a dull appearance and sometimes show flakes of litharge 
on the surface. 

9. Buttons which contain a large amount of platinum flatten out and will not 
blick. They have a steel gray color and dull surface. 

10. A cupel will absorb about its own weight of litharge. 


Second Exercise. Cupellation Assay of Lead Bullion. 


Procedure. Weigh out carefully two or three portions of bullion 
of 4 A. T. each. Wrap each in 10 to 15 grams of silver free lead 
foil so that the whole is very compact, having each piece of lead foil 
of the same size and weight. 

Have a-good fire so that the lead will melt and start to drive with- 





59 
out delay. Use cupels which weigh 35 grams or more and have 
them all in a row with an extra row in front. Drop the assays in 
as quickly as possible and close the door. As soon as the lead starts 
to drive, close the drafts and cool as soon as possible so that feather 
crystal of litharge form on at least the front half of the cupel. Fin- 
_ally open the draft and otherwise increase the temperature for the 
last minute or two of cupellation to drive off the last traces of lead. 
Have some hot cupels in the muffle and as soon as the buttons brighten, 
pull them forward in the muffle to chill and then put a hot cupel over 
them and withdraw both slowly from the muffle. All danger of 
sprouting is over when the inside of the cupel reaches a dull red or 
when the bead has become solid throughout. Remove from the 
furnace to the cupel tray and allow to cool. When the button is 
cold, detach it from the cupel with the button pliers and brush with 
a stiff brush to remove bone-ash, or place it on its side on a clean anvil 
and slightly flatten with a hammer. When free from bone-ash, 
weigh the bead, recording in the note book the weight of gold and 
silver. Then part and weigh the gold, finally report the value of 
gold and silver in oz. per ton. ‘ 

Notes) 1. Have a sheet of clean white paper at hand and when transferring the 
bullion from the scale-pan to the lead foil do it over this so that in case any bullion 


is spilled it will be seen and recovered. Do all of the wrapping and compressing 
over this paper for the same reason. 


2. If the assay is not compact, it may overflow the cupel while melting, or else 
leave small particles on the sides of the cupel, which will not come down into the 
main button. 

Loss of Silver in Cupelling. There is always some loss in cupel- 
lation and this depends on many factors such as the nature and shape 
of the cupel, the temperature of cupellation, the proportion of lead to 
‘silver, the amount and character of impurities, the draft through the 
muffle, etc. Losses may be due to spitting, absorption of bullion 
by the cupel, oxidation and absorption of silver with litharge, and 
volatilization either alone or accompanied by other metals. 

The cupel surface may be regarded as a membrane permeable to 
molten litharge and impermeable to lead. The more nearly the ma- 
terial of the cupel surface approaches this condition the lower the 
losses may be made. Some cupels, particularly some of magnesite, 
- present spots of material which are permeable to lead and consequently 
give a high loss of silver. 

The most important factor relative to cupel loss, however, is the 
temperature. The higher the temperature, the higher the loss, is 
an invariable rule. The increased loss due to higher temperature 
seems to be due mostly to an increased oxidation of the silver and a 


60 


consequent greater absorption loss. The volatilization loss is also 
increased by an increase of temperature. A loss of one per cent 
silver is allowable and the loss usually may be kept close to this figure 
by taking pains to cupel with abundant crystals of litharge. By 


overlooking this matter a loss of 4 or 5 per cent may readily be ob- ' 


tained and this is of course entirely inadmissable. 

The following table taken from Lodge’s Assaying illustrates this 
point and shows the importance of cupelling at the correct tempera- 
ture. The temperature was taken with a Le Chatelier pyrometer, 
the junction being held about ¢’’ above the button. 

TABLE IX. EFFECT OF TEMPERATURE ON LOSS OF SILVER IN 























CUPELLATION. 
Silver 
: Tempera- 
C. P. Sil- | Lead Loss 
ver Mgs. | Grams. pote ae Per ; Remarks. 
nt 1 
cent. 
200 10 700 1.02 Crystals of PbO all around button. 
200 10 775 1.30 Crystals of PbO on cooler side of cupel. 
200 10 850 1.73 No crystals. 
BOC: Sal O 925 3.65 cs z 
200 10 1000 4.88 ‘i : 




















es 


The amount of lead and silver present in any button has a marked 
effect on the percentage loss of silver in cupellation. Rose* in speak- 
ing of cupellation says ‘‘The losses of silver at first are small, so long 
as large quantities of base metals protect it from oxidation.—Later, 
when the percentage of silver is high it is freely oxidized—and the 
oxidation is at its maximum when the silver is practically pure.” 

Keeping the amount of silver constant and varying the lead Lodge 
obtains the results shown in the following table:— 


TABLE X. EFFECT OF LEAD ON LOSS OF SILVER IN CUPELLATION. 

















: Lead Temperature Silver Loss 
Silver Mgs. Grams. Deg. Cent. Per Cent.® 
200 10 685 1.39 
200 15 685 1.38 
200 20 685 1a2 
200 25 685 1.85 








When the quantity of lead remains constant and the silver is varied 
the percentage loss of silver is found to increase as the silver is re- 
duced. The following representative figures taken from Godshall’s 
paper on Silver Losses in Cupellation* show this very clearly. 


1 Average figures. 

2 Trans. Inst. Min. and Met. Vol. 14, p. 420. 

3 Average of two nearest together. : 
aT. A. L. M. E. 26 pp. 473-484 ine. 





61 


TABLE’ XI. EFFECT OF VARYING SILVER ON CUPELLATION LOSSES. 

















Meith or Lead |\1/2 A. 7.1/2 A. T.)1/2 A. T.1/2 A. T.)1/2. A.T.1/2 A. T.]1/2 A. T. 
ea “ Silver|200 Mgs./100 Mgs.| 50 Mgs. | 20 Mgs. | 10 Mgs.| 5 Mgs. | 2 Mgs. 
Silver Loss 1.73% | 2.03% | 2.65% | 2.82% | 344% | 4.46% 6.90% 























Loss of Gold in Cupelling. There is always some loss of gold in 
cupelling but owing to its greater resistance to oxidation this loss is 
smaller than the corresponding silver loss. The following table 
taken from Lodge shows the relation between the loss of gold and the 
temperature of cupellation. 


TABLE XIA. EFFECT OF TEMPERATURE ON LOSS OF GOLD IN 


























CUPELLATION. 
Gold Used Lead Temperature | Gold Loss R i 
Megs. Grams. Deg. Cent. | Per Cent.* aa 
200 10 700 Button Froze. 
200 10 rigs: 0.155 
200 10 850 0.385 
200 | 10 925 0.460 
200 10 1000 1.435 
200 10 1075 2.990 














*Mean of two results nearest together. 


In the case of the gold with temperatures of 1000 degrees and above 
the higher losses seem to be chiefly due to a lessening of the surface 
tension owing to the increased temperature—for on examining the 
cupels with the microscope a large number of minute buttons were 
found all over the inner surface. It would appear that small particles 
of the alloy were left behind to cupel by themselves. 

As in the case of silver the percentage loss of gold is found to in- 
crease as the quantity is reduced. Hillebrand and Allen~ show that 
contrary to the usual opinion, the loss of gold in cupelling is not 
negligible, and is greatly influenced by slight changes in temperatur- 
They found the most exact results to be obtained when feather crys- 
tals of litharge were obtained on the cupels. 








Effect of Silver on the Loss of Gold in Cupelling. Lodge in 
his ‘‘Notes on Assaying” states that the addition of silver in excess 
lessens the loss of gold but gives no figures. Hiullebrand and 
Allen ® state that the loss of gold in cupelling is greater with pure 
gold and alloys poor in silver than with alloys rich in silver. Smiths 


2 Bull. No. 253. U. S. Geol. Survey. p. 20 et seq. 
3 Op cit. 
4 The Behaviour of Tellurium in Assaying. Trans. Inst. Min. and Met 17 p. 472. 


62 


gives the following figures showing the protective action exercised 
by silver on gold during cupellation. 


Per Cent of Total Gold Recovered. 
Tellurium Added Without Tellurium 
Without silver 94.9 98.2 
With silver 97.0 99.5 


Influence of Impurities on the Loss of Precious Metals 
during Cupellation. According to Rose,  tellurium, selenium, 
thallium, bismuth, molybdenum, manganese, copper, vanadium, zinc, 
arsenic, antimony, cadmium, iron and tin, all induce extra losses of 
gold and silver in cupellation and should therefore preferably be 
removed before that stage is reached. 

Of these metals the behavior of tellurium in cupellation will be 
mentioned in the discussion of the assay of telluride ores. Copper 
is perhaps the most common impurity, and on account of the difficulty — 
of removing it completely in scorification or crucible fusions, a knowl- 
edge of its behavior in cupelling is particularly important. Eager 
and Welch’ give the following table showing the effect of copper 
on the loss of silver in cupellation. 


TABLE XIl. EFFECT OF COPPER ON SILVER LOSSES IN CUPELLATION. " 


ee 








Neer | Copper Per Cent Silver Tatioint 

No. sathe Lead Ae ae rcp Lost. Lead to 

; grams. eg: ©: of the man a Ses Copper. 

sata Silver, |) Indiv Wana oe 

1 .20382 10 775 5 1.00 1000 to 1 
2 20256 at a me 1.15 - 
rg .20036 4 a tf 0.93 1.03 as 

4 .20618 «“ 66 10 1.19 500 to 1 
5 .20193 “ “ & ' 1.09 ts 
6 .20118 a : I 06 1.11 te 

7 20146 a“ « 15 1.35 333 to 1 
8 .20138 cf és a 17 
9 20432 tf ee ss 31.15 p ey 

10 20282 e 6 20 31.15 250 to 1 
11 .20100 i tf 1.45 Ee 
5 .20338 4: ad ec 1.46 1.46 . 

13 20224 % es 25 1-058 200 to 1 
14 20496 x ‘ sc 0.95 - 
15 .20420 " - te 1.07 1.02 ? 





1 Jour. Chem. Met. and Min. Soc. of South Africa, Vol. 5, p. 167. 
* Thesis, No. 225, M. I. T. Mining Department. 
8 Disregarded. . . 





j : 


It appears with this lead ratio and temperature that an increase of 
copper from 5 up to 20 per cent of the silver causes a steadily increas- 
ing loss of silver. With 25 per cent of copper the loss is apparently 
less. This was found to be due to the retention of copper in the silver 
button. Comparing the results shown in this table with those shown 
in Table IX, where no copper was used, we find that 5 and 10 per 
cent of copper appears to give lower silver losses than are obtained 
when no copper is present. This may be due to the protective ac- 
tion which copper is known to exert over silver. ' 
Comparing the amounts of lead and copper in 10, 11 and 12 above, 
we find that the ratio of lead to copper should be at least 250 to 1 
to insure the removal of the copper and at least 500 to 1 if the ap- 
parent loss of silver is not to be noticeably increased. 

The effect of copper on the loss of gold is shown in the following 
table :— 


aT 


* TABLE XIII. EFFECT OF COPPER ON GOLD LOSSES IN CUPELLATION. 











Copper Per Cent. Gold 


: 
| 











Ae Pe Lead | Temp. | Per Cent. Lost. we a 
peer. (deg. ©. | of they pcre. 
Gold. Individual} Mean 

iN _.20181 10 TID None 0.15 

2 .20104 _ ie *s 0.16 0.16 

3 .20288 . - 5 0.18 {G00 to | 

4 .20110 . My &, 0.20 es 

5 .20318 i i ef 0.10 Peed! 

(In the following the buttons show a gain in weight.) 

6 .20102 : ts 10 —0.03 500 to 1 

rs .20142 H “ ig —0.03 er 

8 .20138 = e < —0.02 —0.03 7" 

9 .20024 & x. aa Le —0.11 333 to 1 
10 .20060 a peal o —0.26 a 
11 .20048 - ¥ J: —0.18 —0.18 oe 
12 .20100 a ‘S 20m —0.13 250 to 1 
13 .20101 . oH a —0.56* a 
14 .20161 i. e . —0.20 —0.17 e 
15 .20422 cf s 25 —0.29 200 to 1 
16 .20296 3 nM —0.21 és 
7 





cs ie aa Pie 0's? v re0o7 ie “ 





It appears that 5 per cent of copper with this lead ratio has no 
effect on the loss of gold. The gain in the weight of the gold buttons 
with 10 and over per cent of copper shows clearly that copper is 
retained by the gold under these conditions. This was also indicated 
by the color of the gold beads. With a higher cupellation temperature 


1 Rose, Trans. Inst. Min. and Met., Vol. 14, p. 422. 
2 Disregarded. 


64 


the amount of copper retained would doubtless be smaller. It is 
interesting to note that with 10 per cent of copper the amount re- 
tained by the button approximately neutralizes the loss of the gold 
itself. Apparently the ratio of lead to copper should*not be less than 
500 to 1 if the copper is to be completely removed. 


Indications of Metals Present. The behavior of the cupelling 
lead and the appearance of the cupel and bead during and after 
cupellation will often give much valuable information concerning the 
metals present in the bullion. Thus pure lead gives a lemon yellow 
cupel, and bismuth the only other metal which behaves like lead in 
cupelling gives an orange yellow color. Copper gives a green to an 
almost black stain depending on the amount present. If too much 
copper is present the button will freeze, sometimes it will go down 
to a small amount and then flatten out leaving a copper colored 
button. Antimony comes off in the first stages of cupellation giving 
dense fumes and a scoria around the top of the cupel. This scoria 
soon solidifies and expands in so doing. If much antimony is present 
the cupel will be split open by this action allowing the lead to run out 
into the muffle, if present in smaller amounts it may simply crack 
the cupel and leave a ridge of yellow scoria. Arsenic acts much like 
antimony but is not so often carried into the lead button. The scoria 
from arsenical lead is light yellow and the fumes are less noticeable. 
In cupelling buttons containing tin this metal is quickly oxidized 
forming SnOs, which if present in sufficient quantity covers the lead 
with an infusible scoria and stops cupellation. If zzne is present it 
will burn giving a brilliant greenish-white flame. The oxide formed 
condenses on the sides of the cupel and on the lead, and if present in 
quantity will stop cupellation. Alwminum is only slowly oxidized 
at the temperature of cupellation and in cupelling a mixture of the 
two metals it remains behind floating on the lead and is finally left 
on the side of the cupel still in the metallic state. 

Iron, nickel, cobalt and manganese are not easily soluble in 
lead, but if present in any considerable amounts give infusible scoria 
which float on top of the lead and interfere with the cupellation. 
They give no marked characteristic cupel colorings as their oxides 
are only slightly dissolved in the litharge or absorbed by the cupel, 
but usually leave a brown to black stain where the masses of scoria 
come in contact with the cupel. Combinations of the metals influence 
the colors, and copper particularly covers up all of the lighter colors. 

Tellurium gives a pinkish color to the surface of the cupel most of 
which fades away upon cooling. If much tellurium is present it 


Se 


65 


gives a frosted appearance to the bead. Platinum and iridium in 
small amounts act similarly. Using 200 mg. of silver and 10 grams 
of lead the frosting made its appearance when 40 or more mg. of 
tellurium were added. As little as two per cent. of platinum gives 
the silver bead a slightly frosted appearance and 8 or 10 per cent 
gives a very marked rough and frosted look. Buttons which contain 
a large amount of platinum flatten out when near the finishing point 
and refuse to drive leaving a gray, mossy appearing button which 
sticks to the cupel. Such buttons usually retain considerable lead. 
Palladium gives the surface of the lead a raised or embossed appear- 
ance. Ruthenium leaves a black film on the bead and a black scum 
on the cupel.! Osmium behaves somewhat similarly. ' 

Molten gold beads have a beautiful green color and when pure may 
be cooled considerable below the true freezing point. On solidifica- 
tion they ‘‘flash’’ emitting an apple-green colored light. Small 
quantities of iridium, rhodium, osmium, ruthenium, and osmiridium 
prevent this flashing.’ 

In cupelling pure gold or silver with pure lead it is found that the 
cupel will be stained green in the position occupied by the button as 
the last of the lead was going off. The higher the temperature and 
consequently the higher the loss of precious metals, the larger this 
green area becomes. Certain brands of patent cupels give a large 
amount of this green stain, and whenever this is found a serious los; 
of silver is found to have occurred. 


Testing Cupels for Absorption of Silver. An occasional test 
of cupels and especially of each new lot of bone-ash is desirable. 
Select some standard amount of lead and silver and always use the 
same amounts so that results may be comparable. One hundred 
milligrams of silver and 25 grams of lead is a convenient quantity. 
An interesting experiment showing the amount and distribution of 
the silver loss as well as the relative proportion of lead absorbed and 
volatilized may be performed as follows:— 


Procedure. Weigh out carefully on the button balance 100 or 
200 milligrams of C. P. Silver and wrap in exactly 20 grams of C. P. 
lead. Select a hard cupel of bone-ash or magnesia, heat to cupelling 
temperature, cool and weigh. The cupel should be hard enough 
so that there will be no loss by abrasion in handling. 

Cupel carefully with feather litharge crystals. Have a hot cupel 


1 Lodge, Notes on Assaying. 
2 A. J. Van Riemsdijk, Chem. News, Vol. 41, p. 126 and 266. 


66 


at hand to act as a cover and prevent spitting. Clean and weigh the 
silver bead and also weigh the cupel. 

Remove the bone-ash not colored with litharge and grind the re- 
mainder of the cupel to pass a 100 mesh screen. Assay for silver. 


See chapter on crucible assay. Report results as indicated in the 
following example. 





Weight of C. P. silver taken 0.20047 gms. 
Weight of silver after cupellation 0.19725 
Silver lost during cupellation 0.00322 = 1.61% 
Weight of silver found in cupel (silver absorbed) 0.00294 
Weight of silver lost by volatilization 0.00028 
Per cent of lost silver absorbed ay ess 
Per cent of lost silver volatilized 8.9 
Weight of C. P. lead taken 20.00 gms. 
Weight of cupel + PbO 58.71 
Weight of cupel 38.21 
Weight of PbO 20.50 
Weight of lead corresponding 19.04 
Per cent of lead absorbed by cupel 95.20 


Retention of Base Metals. It has already been mentioned that 
a plus error may be incurred by the retention of lead by the silver 
bead. If the bead contains much lead, it will appear dull or slightly 
yellow due to a thin coating of litharge, the bottom where it rests 
against the cupel will be smooth and it will not blick. A button will 
occasionally blick, giving the play of colors, and the “‘flash” even when 
retaining as much as one of two per cent of lead. When the alloy 
contains copper, the silver bead may retain from 2 to 4 per cent of 
copper without showing any unusual symptoms. | 

These retained metals tend to compensate for the absorption loss, 
but are so uncertain that they cannot be counted on to do so. 


Portland Cement and Magnesia Cupels. Cupels of Portland 
cement and calcined magnesia have found favor in some localities. 
The former mostly in the United States and Canada, the latter prin- 
cipally in England and South Africa. Portland cement cupels are 
made from neat cement with from 6 to 10 per cent of water, in the 
usual way. If properly made and handled, they do not crack, and 
they absorb about their own weight of litharge. The silver loss due 
to absorption is greater than for bone-ash. Magnesite cupels are 
mostly factory made, They require a higher muffle temperature 





67 


than bone-ash cupels. They do not crack and absorb about two-thirds 
of their own weight of litharge. An especially high finishing tempera- 
ture is required for magnesite cupels to insure the elimination of the 
last of the lead. The writer has not found them satisfactory for 
silver work. Some makes give high silver losses and with others it 
is found practically impossible to drive off the last of the lead, as 
shown by the cupellation of known amounts of C. P. silver. 

Both of these cupels have one great advantage over bone-ash 
cupels in that when it is necessary to assay the cupel a much better 
slag may be obtained from them than from the very refractory bone- 
ash. Portland cement has the added advantage of cheapness. 


Care of the Muffle. Litharge, being a strong base, quickly 
attacks the material of the muffle. When, therefore, any lead or 
slag is spilled in the muffle, or a fusion is found to have eaten through 
its container, the muffle must be quickly scraped out, and the spot 
well covered with bone-ash. This should be worked around and if 
at all sticky scraped out again and more bone-ash added. 

Muffles last much longer if heated and cooled slowly. When not 
in use the door to the muffle should be kept closed. When through 
with the furnace for the day close all the drafts and the door to the 
muffle and open the top damper so that all parts may covl down 
slowly. 


CHAPTER VI. 


PARTING. 


Parting is the separation of silver from gold by means of acid. In 
gold assaying nitric acid is almost exclusively used, although sul- 
phuric acid is usually employed for parting large lots of bullion. To 
successfully separate silver from gold by the use of nitric acid there 
must be present at least three times as much silver as gold and with 
this ratio the alloy must be in a thin sheet and it requires a long (20 
minutes) continued boiling with acid of 1.26 specific gravity to effect — 
a separation. For parting beads from ore assays it is best to have at 
least six times as much silver as gold present, and for ease of manipula- 
tion we would prefer not to have a much greater ratio of silver to 
gold than this. For with much less silver than this a long continued 
boiling with acid is necessary, while with much more silver than this, 
special precautions have to be taken to prevent the gold from break- 
ing up into small particles which are difficult to manage. The idea 
of parting is to so manipulate that the gold will if possible remain in 
one piece. 

The nitric acid for parting must be free from hydrochloric acid and 
chlorine in order to have no solvent action on the gold and also be- 
cause any chlorides present would precipitate insoluble silver chloride 
on the gold. The acid strength is of great importance and the proper 
strength to be used depends upon the composition of the alloy. The 
higher the ratio of silver in the alloy, the less the acid strength should 
be. 

Great care is necessary in parting to avoid breaking up the gold 
and subsequent loss of some of the small particles, as well as to insure 
complete solution of the silver. 

Different authorities recommend different vessels for parting but 
for ore assays, and especially for beginners in the art, the use of a 
porcelain crucible or capsule is recommended and will be described 
first. Parting in flasks or test tubes with the use of annealing cups 
will also be discussed so that either method may be used. 


Parting in Porcelain Capsules. A glazed porcelain capsule 
12 inches in diameter and 1 inch high is preferable for this work on 








69 


account of its broad flat base, but a small porcelain crucible does very 
well if care is taken not to upset it. Many different strengths of 
acid and other details of manipulation have been recommended but 
the procedure given below is one which has given uniformly satis- 
factory results to the author in his laboratory. The strength of acid 
which may be used depends on the proportion of gold and silver in 
the alloy, the léss the ratio of silver to gold, the stronger the acid may 
’ be without danger of breaking up the gold. It is not intended that 
the method to be described must necessarily be followed in every 
case, but is designed for the safe treatment of buttons having almost 
any proportion of silver to gold, from 3 to 1000 or more parts of silver 
to one of gold. 

Procedure.—Pour into the capsule about 4 inch of dilute nitric 
acid of 1.06 sp. gr. made by diluting 1.42 acid with seven times its 
volume of water. Put on the hot plate and heat until vapor can be 
seen rising from it and then drop in the bead which should be free 
from adhering bone-ash. In case the alloy has only 3 or 4 parts 
of silver to one of gold it must be hammered or rolled out to the 
thickness of an ordinary visiting card, say to 0.01 inch. The bead 
should begin to dissolve at once giving off bubbles of nitrogen oxides. 
If it does not begin to dissolve, add nitric acid 1.26 sp. gr. a few drops 
at a time until action starts. The solution should be kept hot but 
not boiling. The action should be of moderate intensity. Continue 
the heating until action ceases and then decant the solution into a 
clean white evaporating dish in a good light, taking care not to pour 
off any of the gold. Then add a few cubic centimeters of 1.26 
sp. gr. acid, made by diluting strong nitric acid 1.42 sp. gr. with 
£ its volume of water, and heat almost to boiling (90° C.) for from two 
to ten minutes. Decant this solution and then wash three times with 
warm distilled water, decanting as completely as possible after each 
washing. Apply the stream of water from the wash bottle tangen- 
tially to the sides of the capsule, rotating it meanwhile to prevent di- 
rect impact of the stream on the gold. After the final washing 
manipulate the particles of gold so as to bring them together, decant 
off the last drops of water as completely as possible and set the cup 
in a warm plate to dry the gold, but avoid too high a temperature 
as the sputtering of the last drop of water would tend to break up 
and possibly throw out the gold. Finally ‘‘anneal”’ the gold by put- 
ting the cup in the muffle or over the open flame until the bottom is 
bright red, when the gold will change from its black amorphous 
condition to the true yellow color of pure gold. It is now ready to 
cool and weigh. To transfer the gold from the cup to the scale-pan, 


70 


bring the scale-pan to the front part of the balance. Gradually 
invert the cup over the pan, tapping it meanwhile with a pencil, when 
the gold will usually slide out without difficulty. If any small particles 
stick to the cup they may be detached by touching them gently with 
the point of the forceps or a small camels hair-brush. 

The gold should be pure yellow throughout and may be compared 


with parted gold of known purity. If it is lighter colored than pure | 


gold it is probable that all of the silver has not been dissolved. If 
it is dark in spots or if the cup is stained, it indicates incomplete re- 
moval of the silver nitrate. The ‘annealing’? causes the gold to 
stick together making it easier to handle, tends to burn out any 
specks of organic matter which may have fallen into the cup, allows 
us to observe the color of the parted gold and to determine its purity 
in that way and to distinguish and separate any specks of foreign 
matter such as fire brick, coke dust etc. which may have found their 
way into the cup. The ‘‘annealing’’ at a red heat is also necessary 
in order that the gold may contract and lose most of its porosity, 
since otherwise it would condense a considerable quantity of gas 
during weighing. 

After the silver has been dissolved from a doré alloy by the acid, 
the gold remains as a porous mass which is more compact the larger 
the proportion of gold the alloy contained, the thicker the alloy and 
the less the mechanical disturbance of the bead during solution. In 
treating a bead which is near the limiting ratio of silver to gold it is 
sometimes difficult to determine whether or not it is parted. This 
may be ascertained by touching it with a glass rod drawn down to 
a rather small diameter, (approximately 1/32 inch). If it feels soft 
throughout and can be broken up it is practically parted but it should 
be heated almost to boiling with 1.26 sp. gr. acid for at least ten min- 
utes to insure dissolving the last of the silver. Such a mass of parted 
gold will require a longer and more careful washing, for on account 
of its density a longer time is required for the silver nitrate to diffuse 
through its minute pores. In parting the ordinary bead containing 
10, 20 or more times as much silver as gold, it is easily seen when 
parting is complete by the considerable shrinking of the mass. 


Inquartation. When the bead contains too little silver to part 
(less than three parts of silver to one of gold), it is necessary to alloy 
it with more silver in order to get the gold in a pure state. To do 
this, wash and dry the bead and wrap it up in say six times its weight 
of silver and fuse it on a piece of charcoal by means of a blow pipe, 
or better wrap the whole in 4 or 5 grams of sheet lead and cupel. 


as 
7 a ee 





rea 


The term inquartation originated from the custom of the old assayers 
of adding three-quarters of silver to one-quarter of gold. 

Many assayers when working for both gold and silver and suspect- 
ing an ore to be deficient in silver, add silver to the crucible or to the 
lead button before cupelling, part directly and then run separate 
scorification assays to determine the silver in the ore. 


Preparing Large Beads for Parting: Large beads especially 
those which approach the maximum ratio of 25 per cent gold must be 
flattened on an anvil and rolled out to a thickness of about 0.01 
inch before parting. During this process the alloy will require fre- 
quent annealing to prevent it from cracking. It should finally be 
rolled up into a little “‘cornet’’ before parting. (See assay of gold 
bullion.) 


Notes: 1. The nitric acid solution should be hot before dropping in the bead as 
in cold acid the gold tends to break up into extremely fine particles. 


2. The violent mechanical disturbance due to boiling or to too rapid solution may 
cause the gold to break up causing difficulty or actual loss in washing and subse- 
quent handling. 


3. If there remains only a few tenths of a milligram of porous gold the 10 minutes 
heating with 1.26 sp. gr. acid is unnecessary. 


4. Strong nitric acid (1.46 sp. gr.) should not be used at any time as gold is slightly 
dissolved by it. 


5. If in doubt at any time as to the purity of your parted gold, wrap it up in six 
times its weight of silver foiland carefully cupel with lead, finally re-part and weigh. 


6. If a small particle of gold is seen floating on the surface of the liquid, it may be 
made to sink by touching it with a glass rod. 


7. The black stain occurring in parting cups after heating is due to metallic silver 
reduced from silver nitrate by the heat, showing insufficient washing. 


Parting in Flasks etc. Parting in flasks, test tubes, etc. is, up 
to the completion of the washing of the gold, exactly similar to that 
in porcelain capsules. From this point on, however, the manipula- 
tions are different, as the annealing is not done in the same vessel, 
but in an annealing cup. The annealing cup is a small unglazed 
crucible made of fire clay and very smooth on the inside. 

Procedure:—After washing the gold, fill the flask or test tube with 
distilled water, invert over it an annealing cup and then quickly 
invert the two so that the gold may fall into the cup. This operation 
should be done in a good light and preferably against a white back- 
ground. Tap the flask if necessary to dislodge any gold which has 
caught on the side and after all the gold has settled, raise the flask 
slowly until its lip is level with the top of the annealing cup. Now 
when all the gold is at the bottom of the cup, slip the flask quickly 
from the cup and invert it. Drain the water from the cup, cover it 


72 


and set it on the hot plate to dry. When fully dry, it is ready to 
anneal and weigh. Examine the flask once more to make sure that 
no gold has been left in it. 

This method of parting has the advantage that the acid may be 
boiled if necessary with less danger of its boiling over and causing 
loss of fine gold. It is well suited for the parting of large beads where 
the porcelain cup would not contain enough acid to dissolve all of 
the silver, and also to the parting of alloys where the ratio of silver to 
gold is only 2 or 3 to one and which would therefore require a long 
continued heating at or near boiling temperature. The methods is 
therefore recommended for use in the assay of gold bullion. The 
clay cups have the advantage of porosity so that they can absorb the 
last drops of water and give it off again slowly, thus preventing spat- 
tering if they are set on a hot iron plate to dry. They also stand 
sudden changes of temperature somewhat better than the glazed 
porcelain cups. 

This method has the disadvantage that if all the parted gold does 
not remain in one piece, there is greater danger of loss because the fine 
gold settles with difficulty and because it cannot be watched so well 
through all stages of the process. There is also danger of small 
particles of the cup and especially the cover being broken off and 
mixed with the gold. 





CHAPTER VII. 


THE SCORIFICATION ASSAY. 


The scorification assay is the simplest method for the determination 
of gold and silver in ores and furnace products. It consists simply of 
an oxidizing muffle fusion of the ore with granulated lead and borax 
glass. The lead oxide formed combines with the silica of the ore and 
also to a certain extent dissolves the oxides of the other metals. The 
only reagents used other than lead are borax glass and occasionally 
powdered silica, which aid in the slagging of the basic oxides. 

_ The scorifier is a shallow, circular fire-clay dish 2 or 3 inches in 
diameter. The sizes most commonly used are 24, 23 and 3 inches in 
diameter. 

The amount of ore used varies from 0.05 A. T. to 0.25 A. T., the 
amount most commonly used being 0.10 A. T. With this is used 
from 30 to 70 grams of test lead and from 1 to 5 grams of borax 
glass, depending on the amount of base metal impurities present. 
With nearly pure galena, or a mixture of galena and silica, a charge 
of 30 to 35 grams test lead and 1 gram of borax glass will suffice for 
0.10 A. T. of ore, but when the ore contains nickel, copper, cobalt, 
arsenic, antimony, zine, iron, tin etc. a larger and larger amount of 
lead and borax glass must be used dependent upon the relative slagabil- 
ity of the metals and the solubility of their oxides in the slags formed. 
Of the above, nickel and copper especially are very difficultly oxidized 
and when much of these are present in the ore the lead button from 
the first scorification will have to be rescorified once or twice with 
added lead. Iron, on the other hand is comparatively readily oxidized, 
and aside from adding an extra amount of lead and borax glass to make 
a fluid, slag the ore is as readily assayed as galena. Lime, zinc and 
antimony especially require large amounts of borax glass to convert 
their refractory oxides into a fusible slag. 


Solubility of Metallic Oxides in Litharge. Litharge, although 
a strong base has the power of holding in igneous solution certain 
quantities of other metallic oxides. This has an important bearing 
on the ease or difficulty with which various metals may be slagged in 
scorification. According to Berthier and Percy the solubilities of 


74 


the various metallic oxides in litharge are as shown in the following 
table: 
One part of Cu2ed CuO ZnO FeO; MnO Sn0O, Ti0s 
Requires parts of PbO 1.5 1.8 8 10 10 12 8 
Antimony trioxide (Sb2OQ3) dissolves in litharge in all proportions. 
Heat of Formation of Metal Oxides. Another important 
factor having to do with the elimination of impurities by scorification 
is the relative heat of formation of the various metal oxides. Having 
a mixture of various metal sulphides and assuming for a moment 
the ignition temperature to be the same for all, that reaction in which 
is evolved the greatest amount of heat would naturally proceed at 
the fastest rate. The heat of combination of various metals each 
with 16 grams of oxygen is shown in the following table. This basis 
being used on the assumption that the amount of oxygen is limited. 
TABLE XIV. HEAT OF FORMATION OF METALLIC OXIDES. 














Reaction. Heat of Comb. with 16g0. Reaction. Heat of Comb. with 16g0. 








Zine to ZuO 870 1378 


1378 Bismuth to BisOs a = 459 

Tin to SnO2 v3 = O89 Copper to CuO 74 
657 3 692 

whee : ar 638 Sulphur to SOe “9 = 346 
Nickel to NiO 579 : : 386 

Lead to PbO 503 Tellurium to TeOz “9 = 198 

2312 Bia 
Antimony to Sb20s AON = 462 Silv a Ag:O =158 70 
2194 Gold to AusO3 3 = a8 
Arsenic to As2Os oi 439 














Ignition Temperature of Metallic Sulphides. The ignition 
temperature of the metallic sulphides may also be used as an indica- 
tion of the order in which the various metals will be eliminated in 
scorification. 


TABLE XV. IGNITION TEMPERATURES OF METALLIC SULPHIDES ~ 
WHEN HEATED IN AIR. 














“1 i Ignition : : Ignitoin 
Material Formula | Temp.° C.: | Material Formula Temp.” Cc. 
Stibnite Sb2O03 | 290-340 Galena ! PbS 554-847 
Pyrite FeSe 325-427 Millerite NiS 573-616 
Pyrrhotite |FexSx+1 _ 4380-590 Argentite | AgeS 605-873 
Chalcocite |Cu2S 430-679 Sphalerite | ZnS 647-810 








Nickel is by far the hardest metal to eliminate in scorification and 
none of the above figures exactly explain this. It is very much harder — 
to slag than either copper or cobalt. 

' In Oxygen. 


75 


Notes. 1. Litharge being a strong base has a great affinity for the silica of the 


__scorifier and especially when mixed with copper oxide it attacks it readily. When 


scorifying matte and copper bullion it is often necessary to add powdered silica to 
the charge to prevent a hole being eaten through the scorifier. 


2. Some assayers add litharge to the scorification charge especially with pyritic- 
ores. On heating, the litharge is reduced to metallic lead, the sulphur of the pyrite 
being oxidized. 


Scorification Assay of Silver Ore. 


Procedure:—Empty the bottle or envelope of ore onto a sheet of 
glazed paper or oil cloth and mix thoroughly by rolling. 

Take three scorifiers, 245, 2%, and 3 inches in diameter respectively. 
Weigh out on the flux balance three portions of granulated lead 35, 
45, and 55 grams respectively. Divide each lot of lead approximately 
in halves, transfer one half of each to the corresponding scorifier 
and reserve the remaining portions. Weigh out three portions of 
exactly 0.1 A. T. of ore on the pulp balance and place on top of the 
lead in the scorifiers. Mix thoroughly with the spatula and cover 
with the remaining portions of lead. Scatter one or two grams of 
borax glass on top of the lead. The scorifiers are now ready for the 
muffle, which should be bright red or yellow before the charges are 
put in and this temperature should be maintained during the roasting 
period. 
Fusion Period. Place the scorifiers well back in the muffle, close 
the door and allow the contents to become thoroughly fused. 
Roasting Period. When thoroughly fused, open the door to admit 
air to oxidize the ore and lead. If the ore contains sulphides these 


will now be seen floating on the top of the molten lead. The sulphur 


from these is burned going off as SO, and the base metals are oxidized 
and slagged. The precious metals remain unoxidized and are taken 
up by the lead bath. These patches of ore grow smaller and soon 
disappear, after which the surface of the melt becomes smooth, con- 
sisting of a bath of molten lead surrounded by a ring of slag. 

The vapor rising from the assays will often indicate the character 
of the ore. Sulphur gives clear gray fumes, arsenic grayish white and 
antimony reddish. Zinc vapor is blackish and the zinc itself may be 
seen burning with a bright white flame. 

Scorification Period. The lead continues to oxidize and the ring 
of slag around the circumference of the scorifier becomes larger as 
more of the lead is oxidized. Finally the whole of the lead is covered 
with slag and the scorification is finished. The ore should be com- 
pletely decomposed and practically all of the gold and silver should 
be alloyed with the metallic lead. 

Liquifaction Period. Close the door of the muffle and increase the 





76 


heat for a few minutes to make the slag thoroughly liquid and to 
insure a clean pour. Then pour the contents of the scorifiers into a 
dry, warm, scorifier mold which has been previously coated with 
chalk or iron oxide. Pour carefully into the center of the mold or 
else the lead is likely to spatter and may not all come together in one 
piece. The inside surface of the scorifiers should be smooth and glassy 
showing no lumps of ore or undecomposed material. 

When cold, separate the lead from the slag, hammer into the form 
of a cube and weigh to the nearest gram on the flux balance. Exam- 
ine the slag and sides of the mold carefully for shots of lead and if 
any are found add them to the main button. 

If the lead is soft and malleable, and the color of the scorifier does 
not indicate the presence of large amounts of copper, nickel or cobalt, 
the button is ready for cupellation. If it is hard or brittle it may con- 
tain impurities which must be removed by rescorifying with an addi- 
tional amount of granulated lead. 

Finally cupel and weigh the resultant silver or doré beads. Re- 
port in your notes the weight of ore and reagents used, the weight 
of lead button obtained as well as the weight and assay in oz. per 
ton of gold and silver. Note also the time of scorification and cupel- 
lation and describe the appearance of the scorifier and cupel. 


Notes. 1. The ore must be so fine that a sample of 1/10 A..T. of it can be taken 
which will truly represent the whole; 100 mesh may be fine enough for some ores, 
170 mesh may be necessary for some others. 

2. In weighing out the ore, spread the sample which has been thoroughly mixed, 
into a thin sheet on the glazed paper at one side of the pulp balance. Place the 
weight on the right-hand pan and the ore on the left-hand pan. With the spatula 
mark the ore off into squares 1 inch or so on a side and then take a small portion 
from every square for the sample, being sure to take a section from top to bottom 
of the ore. During this first sampling the scale-pan should be held over the paper 
in one hand and the spatula in the other. When what is judged to be the right 
amount of ore is obtained the pan is put back on the balance and the hand with 
which it was held is used to turn the balance key. 

The balance should be turned out of action each time ore is put on or taken off 
the scale-pan and the pointer need move only 1 or 2 divisions to indicate whether 
too much or too little ore is on the pan. To obtain the final balance, have a little 
too much ore in the pan, take off enough on the point of the spatule so that the con- 
dition of balance is reversed. With the balance key lift the beam only sufficient 
to allow the pointer to swing one or two divisions to the left of the center and then 
hold the key in this position. Hold the spatula over the pan and by tapping it 
gently with the first finger allow the ore to slide off onto the scale-pan a few grains 
at a time, until the balance is restored and the needle swings over to the center. 
By repeating this process, rejecting the ore retained on the spatula each time, an 
exact weight can soon b2 obtained. 

3. If the contents of the scorifiers do not become thoroughly liquid and show a 
smooth surface of slag after 10 or 15 minutes, the assays require either more heat, 
more borax glass or more lead. 

4. The lead button should weigh from 12 to 20 grams. If it is much smaller than 
this there is danger of a loss of silver due to oxidation, especially when the ore is rich. 
If the button is too large it may be rescorified in a new scorifier to the size desired. 

5. The size of scorifier to be used depends upon the amount of ore, lead, borax 
glass and silica used, and should be such as to give a button of approximately 15 to 





77 


18 grams. If a large scorifier is used with a small amount of lead the resulting lead 
button will be very small and a high loss of silver will result. Again, the larger the 
amount of borax glass that is used the more slag there will be and the sooner the lead 
will be covered over. 

6. Hard buttons may be due to copper, antimony or in fact almost any metal 
alloyed with the lead. Brittle buttons may be due to one of many alloyed metals, 
or to the presence of sulphur or lead oxide. 

7. Ores containing pyrite require a higher temperature during the roasting period 
than those containing galena. 

8. The white patches occasionally found in the slag are made up mostly of lead 
sulphate which is formed when the scorification temperature is low. 

9. Instead of weighing the granulated lead, it may be measured with sufficient 
accuracy by the use of a shot measure or small crucible. The borax glass may also 
be measured. 


Chemical Reactions in Scorification. 
Simple Oxidation. At first, after the lead is melted and the air 
is admitted to the muffle, the lead begins to oxidize to PbO and this 
oxidation continues through the whole scorification period. 


Roasting Reactions. The sulphides in the ore are roasted as 
indicated by the following reactions :— 
PbS + 80 = PbO + SO, 
2PbS + 70 = PbO + PbSO, + SO 
ZnS + 30 = ZnO + SO. 
SbeSs + 90 = Sb2O3 + 3802 
Iron pyrite breaks up on heating as follows:— 
FeS. + heat = FeS + 8 
after which the sulphur is oxidized and the iron sulphide roasts. 


Slag Forming Reactions. The litharge formed combines with 
the siliceous gangue of the ore forming silicates. The borax also 
combines with the various metal oxides forming borates:— 

4PbO + SiO, = Ph.SiO¢ sub-silicate 
FeO + NasB,O; = NaeFeB:Os 


Reactions Between Sulphides and Oxides. After enough PbO 
has been formed to slag the siliceous gangue, the litharge which is 
formed reacts on the partially decomposed sulphides aiding in the 
elimination of sulphur, thus: 

PbS + 2PbO = 3Pb + SO 
ZnS + 3PbO = 3Pb. + ZnO + SO; 
AgeS + 2PbO = 2Pb.Ag + SO 
Part of the arsenic Balntiltaas and part goes into the slag. 

If CuS were present in the ore, part of it would be oxidized to 
CuO and then the cuprous, sulphide and the cupric oxide would tend 
to react as follows: : 

Cues + 2CuO = 4Cu + SO» 


78 


A similar reaction between the litharge and the cuprous sulphide 

would probably take place as follows: 

Cues8 + 2PbO = 2CuPb + SO, 
The copper thus reduced and alloyed with the lead requires a pro- 
longed scorification to remove. The two last reactions are more 
pronounced at high temperatures, so that for the elimination of copper 
in the scorification assay it is evident that a low muffle temperature 
should be maintained. 

The color of the thin coating of slag on the scorifier is an indication 
of the amount and kind of metal originally present in the ore, and 
taken in connection with the mineralogical examination of the ore it 
gives a very good approximation as to its composition. 

Copper gives a light or dark green depending on the amount pres- 
ent. If there is much iron in the ore this color may be wholly or in 
part obscured by the black of the iron oxide. The iron is practically 
wholly removed in the first scorification so that in assaying a copper 
matte the first scorifier may appear black while the second one will 
be green. The green color is said to be due to a mixture of blue 
cupric silicate and yellow lead silicate. 

Iron. A large amount of iron makes the scorifier black, from which 
the color ranges from a deep red through various shades of brown to 
a yellow brown. 

Lead in the absence of other metals makes the scorifier lemon- 
yellow to a very pale yellow. 

Cobalt gives a beautiful blue if other metals do not interfere. 

Nickel colors the scorifier brown to black depending on the amount 
present. When much nickel is present the cupel becomes covered 
with a thick film of green nickel oxide. 

Manganese colors the scorifier brownish black to a beautiful wine 
color. 

Arsenic and antimony if present in large amount, will leave crusts 
on the inner surface of the scorifier even if much borax glass is used.’ 
In the absence of other metals these scoria will be yellow in color. 

If a scorifier is colored dark green, indicating much copper, dark 
blue, indicating much cobalt, or black with infusible scoria, indicating 
nickel the button should be seorified again with more lead. 


Rescorifying Buttons: When rescorifying to remove copper or 
other impurities, add sufficient lead to bring the total amount of lead 
up to 50 grams and scorify at a low temperature using a 3’’ scorifier. 
Place the scorifier in the muffle, heat to scorifying temperature (to 

1 Lodge. 





79 


remove moisture) and then drop in the lead. Sometimes a button 
requires as many as three scorifications before it is Ske ea hy HN 
to cupel. 

Buttons weighing over 30 grams should be wentital 66° 12" or TS 
grams before being cupelled, as the loss of silver should be less by the 
combined method than by direct cupellation. ; 


Spitting of Scorifiers. Occasionally small particles of lead are 
seen being projected out of the scorifier. This is due to decrepita- 
tion of the ore or to the action of some gas given off by the ore or scori- 
fier itself. If the particles of lead do not all fall back into the scori- 
fier a loss of precious metal will result. The direct cause may be 
found in some of the following and a proper remedy applied: 

1. Dampness of scorifier. 

2. Presence of carbonates in clay from which scorifier was made. 

3. Imperfect mixing of charge, resulting in ore being left on the 
bottom of the scorifier and covered with lead. 

4. Too high a temperature at the start, resulting in too rapid oxida- 
tion of sulphides, evolution of CO: or violent decrepitation. | 

5. Admittance of air into the muffle too soon, resulting in too rapid 
oxidation. Especially to be avoided in the case of ores or products 
carrying zinc. 7 

6. Character of the ore itself. Ores containing carbonates etc. 
are not suited for scorification. 


Assaying Granulated Lead. Almost all assay reagents contain 
traces of gold and silver, but the lead and litharge especially are most 
likely to contain these metals in appreciable amounts. Each new lot 
of granulated lead which is obtained should be sampled and assayed 
before it is used, and in case any silver or gold is found a strict account 
must be kept of the lead used in each assay and a correction for its 
precious metal contents made. 

Procedure: Scorify 2 or 3 portions of 120 grams each in 34 or 4’’ 
scorifiers. If necessary rescorify until the buttons are reduced to 
15 or 20 grams. . Cupel, weigh and part. This correction must be 
made even if extremely small, as any error thus introduced would be 
multiplied by 10 in reporting the results in oz. per ton. 


_ Scorification Assay for Gold. The silver in an ore can be de- 
termined with a sufficient degree of accuracy by taking 1/10 A. T. 
for each assay, since we may thus determine the contents of the ore 
to 1/10 of an ounce, or its value to 5 or 6centsaton. When however 


80 


we determine gold to 1/10 of an ounce per ton by this same method, 
we have determined its value to only 2.00 per ton, which is not 
sufficiently accurate for any but very high grade ores. For this rea-. 
son the scorification assay is not usually chosen for gold ores unless 
they contain impurities which interfere decidedly with the crucible 
assay. (See scorification assay of copper matte and bullion.) 


Scorification Assay of Copper Matte. 


Procedure: Take three portions of 1/10 A. T. of matte, and mix 
with 30 grams of granulated lead and 1 gram powdered silica in a 
3-inch scorifier and cover with 30 grams more of lead. Put 4 gram 
of borax glass on top. Scorify hot at first and then at a low tempera- 
ture to facilitate slagging the copper. 

When the lead eye covers, pour as usual and separate the lead from 
the slag. Weigh each button and add sufficient granulated lead to 
bring the total weight to 50 grams and drop into three new scorifiers 
which have been heated in the muffle. Add about 1 gram of silica 
and scorify at a low temperature. 

Repeat this second scorification if necessary until the cool scorifiers 
are light green. Cupel as usual. The color of the cupel should be 
greenish and not black. The latter color indicates insufficient 
scorification. 

Weigh the combined silver and gold and part, weighing the gold. 

See special methods of assay for details of this and other methods 
of assay of copper-bearing material. 


Notes. 1. For matte containing not more than 30 per cent of eopper two scori- 
fications are sufficient. 

2. This method gives rather high slag and cupel losses and for exact work the 
slags and cupels are re-assayed and a correction made for their silver and gold con- 
tents. 

3. The final silver beads will often contain from 2 to 4 per cent of copper. 

4. When accurate results in gold are desired as many as 10 portions of 1/10 A. T. 
each of matte are scorified and the buttons combined for parting and weighing. 


The scorification process is particularly suited to sulphides, arsen- 
ides and antimonides of the difficultly oxidizable base metals, par- 
ticularly nickel, copper and cobalt. It is used in many localities 
for the silver in all sulphide ores, as well as for copper and nickel mattes. 
When gold is to be determined some other method is generally used. 
The scorification assay is not commonly used for the assay of siliceous 
ores. | 

The following changes have been found generally satisfactory :— 





81 


TABLE XVI. SCORIFICATION CHARGES FOR DIFFERENT ORES. 























Charge 
’ ane 
.| Granu- Scori- | High at 
Ore Ore lated | Borax fier First 
ley Lead Glass | Silica then 
ems. gms. | gms. 
Galena 0.1 30 4-] ~ 24’’ Low 
+ Galena 3 Silica 0.1 35 4-1 ~ 24’’ Low 
Low grade galena 0.2 45 4-] - 2i”’ Low 
Pyrite 0.1 50 2-3 — 23"' Medium 
3 Pyrite 4 Silica 0.1 45 1-2 _ 24’’ Medium 
Stibnite 0.1 50-60 1-2 - 23-3"’ High 
Sphalerite 0.1 60 3-5 1-2 oy High 
Arsenical ores 0.1 45-60 | -1-2 — 23-3'' High 
Cobalt ore 0.1 60 3 — ais High 
Nickel ore 0.05-0.1 60 3 ~ au High 
Chalcopyrite 0.1 60 1-2 1 iad Low 
Tin ores 0.1 60-70 2-3 1 3-34'’ High 
Lead matte 0.1 50 3 - 23/’ Low 
Copper matte 0.1 60 1 1 oe Low 




















Losses in Scorification. Losses in scorification may be due to 
“spitting,” volatilization, oxidation and slagging as well as losses of 
shots of alloy. Some loss due to oxidation and slagging is unavoidable, 
but it should be low. If there is any decided loss by volatilization it 
shows that the process is unsuited to the ore. 

The tendency of scorification assays to “‘spit’’-is one of the most 
serious objections to the process. Ores which decrepitate or contain 
volatile constituents such as COs, H2O, ete. (CaCO;, CaSO,.2H20) 
are unsuited to the process and should be assayed by crucible methods. 
_ Very often a preliminary glazing of the scorifier with a mixture of 
sodium carbonate and borax-glass will prevent spitting. The scori- 
fiers should always be kept in a warm, dry place. 

Losses of alloy due to failure of all the lead to collect in one piece 
may be due to careless pouring, by which some of the lead may splash 
on the side of the mold and solidify there, or on account of poor 
slag, or a cold pour, some shots of alloy may be left in the scorifier 
or scattered through the slag in the mold. 

As scorification is an oxidizing process it is only reasonable to 
expect some loss due to oxidation of the precious metals, and this will 
~ naturally be greater the longer the scorification is continued and the 
- more intense the oxidizing action. Silver is more easily oxidized 
~ than gold, therefore we should expect a much greater loss of silver than 
of gold. To keep this loss at a minimum let the liquifaction period 
be thorough. The molten lead tends to reduce and collect some of 





82 








the silver previously slagged. Some assayers recone 
a small amount (about 0.2 grams) of charcoal over the slag ] e 
scorifier just before pouring, with the idea of reducing some lead fron E: 
the slag and thus collecting most of the oxidized silver by the rai n 
of lead shot thus induced. English authorities almost invariably | 
recommend this practice which they term ‘ ‘cleaning the slag.” 





CHAPTER VIII. 


THE CRUCIBLE ASSAY. 


Theory of the Crucible Assay. The majority of ores are in- 
fusible, or nearly so, by themselves, but if pulverized and mixed with 
suitable reagents in proper proportion the mass will readily fuse at 
an easily attained temperature. The finer the ore is crushed, the 
better and more uniform are the results usually obtained. We as- 
sume in considering a crucible assay that there is such a thorough 
mixture of ore and fluxes that each particle of ore is in contact with 
one or more particles of litharge and reducing agent. As the tempera- 
ture of the mass is gradually raised, part of the litharge is reduced to 

lead (commencing at 500° to 550° C.) by the carbon of the charge and 
- these reduced shots of lead, alloy and take up the gold and silver from 
the surrounding particles of ore, so far at least as the precious metals 
are free to alloy. 

At about this same temperature, 560° C., the borax of the charge 
begins to melt and to form fusible compounds with some of the bases 
of the flux and ore charge. At 625° C. lead oxide and silica commence 
to combine and from this point the slag begins to form rapidly. The 
conditions should be such that the slag remains viscous until the ore 
particles are thoroughly decomposed and every particle of gold and 
silver has been taken up by the adjacent suspended globules of lead. 
After this point has been passed, the temperature may be raised until 
the slag is thoroughly fluid, when the lead particles combine and fall- 
ing through the slag form a button in the bottom of the crucible in 
which practically all of the precious metals originally present in the 
ore are concentrated. 

To make an intelligent crucible assay it is necessary to know the 
mineral character of the ore, for a siliceous ore requires a different 
treatment from one which is mostly limestone and a sulphide requires 
to be treated differently from an oxide. For the purpose of the as- 
sayer, ores should be considered from two standpoints, first according 
to the character and quantity of their slag forming constituents, 
and second according as they are oxidizing, neutral or reducing in 
the crucible fusion with lead and lead oxide. 


84 


Ores Classified According to Slag Forming Constituents. 
The principal slag forming constituents of ores and gangue minerals 
arranged approximately in the order of their occurrence in the earth’s 
crust are as follows:— 


Silica S10. Acid 
Alumina Al.O3_ 
Ferrous-oxide FeO 
Manganous-oxide MnO 
Calcium-oxide CaO 
Magnesium oxide MgO |, 
Sodium oxide Na2O Bh: 
Potassium oxide KO 

Zine oxide ZnO 

Lead oxide- PbO 
Cuprous oxide Ci. 


These oxides with the exception of those of sodium, potassium and 
lead are infusible at the temperature of the assay furnace. To get 
them into the molten condition we add fluxes. According to Percy, 
‘a flux is something which if added to a substance infusible or diffi- 
cultly fusible by itself will cause it to fuse.”’ 

All of the common assay fluxes with the exception of silica are 
readily fusible by themselves. In general it may be said that to 
flux the acid, silica, it is necessary to add bases, and to flux any of 
the basic oxides acids must be added. To flux alumina it is best 
to add both acids and bases, thus making what is termed a double . 
silicate. 


Ores Classified According to Oxidizing or Reducing Charac- 
ter. According to their oxidizing or reducing character in the crucible 
assay ores may be divided into three classes as follows: 

Class 1. Neutral Ores. Siliceous, oxide and carbonate ores or 
ores containing no sulphides, arsenides, antimonides, tellurides, 
etc., i.e. ores having no reducing or oxidizing power. 

Class 2. Ores Having a Reducing Power. Ores containing sul- 
phides, arsenides, antimonides, tellurides etc. or containing carbon- 
aceous matter, or ores which decompose litharge with a reduction of 
lead in the crucible fusion. 

Class 3. Ores Having an Oxidizing Power. Ores containing fer- 
ric oxide, manganese dioxide etc. or ores which when fused with fluxes 
oxidize lead or reducing agents. Ores with any considerable oxidiz- 
ing power are comparatively rare. 


85 


Determining the Character of a Sample. The mineral char- 
acter of an ore can be most readily determined when the ore is in the 
coarse condition. As however a large proportion of the samples re- 
ceived by the assayer are already pulverized, it becomes necessary 
for him to be able to form a close estimate of their composition in 
this condition. This may be best accomplished by washing a small 
sample on a vanning placque or shovel. 

Vanning. Place one or two grams of the ore on the vanning shovel, 


 eover it with water and allow it to stand until the ore is thoroughly 


wet, shake violently in a horizontal plane until the fine slimes are in 
suspension and all lumps are broken up. Allow it to settle a moment, 
decant some of the water if necessary and then separate the ore ac- 
cording to the specific gravity of its different minerals by a combined 
washing and shaking. The water should be made to flow over the 
ore in one direction only and the velocity of the shaking motion should 
be accelerated in a direction opposite to the flow of the water. The 
shaking tends to stratify the ore, heaviest next the pan, lightest 
on top, while the water tends to wash everything downward, the 
material on top most because of its position, and also because of its 
lesser specific gravity. Finally if there are a number of minerals 
present, they should appear spread out in fan shape in order of their 
specific gravity, for instance, galena, pyrite, sphalerite and quartz. 


Crucible Slags. The slags obtained in the crucible assay may be 
regarded as silicates and borates of the metallic oxides. The acid 
constituents of rocks other than silica so seldom play an important 
part in the formation of slags that they may be omitted at least from 
a preliminary discussion of the subject. 

A slag suitable for assay purposes should have the following proper- 
ties :— 

1. A comparatively low formation temperature readily attainable 
in assay furnaces. 

2. It should be pasty at and near its formation temperature, to 
hold up the particles of reduced lead until the precious metals are 
liberated from their mechanical or chemical bonds and are free to 
alloy with the lead. 

3. It should be thin and fluid when heated somewhat above its 
melting point, so that shots of lead may settle through it readily. 


4. It should have a low capacity for gold and silver, and should 


allow a complete decomposition of the ore by the fluxes. 





5. It should not attack the material of the crucible too violently. 


86 


6. Its specific gravity should be low, to allow of a good separation 
of lead and slag. 


7. When cold, it should separate readily from the lead. 


Classification of Silicates. 


Silicates are classified according to 


the proportion of the oxygen in the acid to oxygen in the base. Thus 
a mono-silicate has the same amount of oxygen in the acid as in the 
base. A bi-silicate has twice as much oxygen in the acid as in the 


base and so on. 





The silicates which have been found to behave satisfactorily as | 


assay slags lie within the following limits:— 


TABLE XVII. CLASSIFICATION OF SILICATES. 

















Nae Oxygen Ratio Formula. R = Bivalent 

: Acid to Base Base 
Sub-silicate it ete 4RO.Si02 
Mono-silicate 1to4 2RO.Si02 
Sesqui-silicate 13 to 1 4RO.3S102 
Bi-silicate 2 to 1 RO.Si02 
Tri-silicate 3 to 1 2RO.38102 











The formation temperature and melting point of the different sili- 
cates depends not only on the relation of the silica to base, but also 
on the nature of the bases present. ‘Thus we say that within the 
above range the silicates of lead and the alkalies are all readily fusible, 
the iron and manganese silicates are difficultly fusible and the silicates 
of calcium, magnesium and aluminum are infusible at the temperature 
of the assay furnace. Note that so far we are referring to the indi- 
vidual silicates of the different bases and not to mixtures of the same. 

Of these slags the bi- and the tri-silicates have but little effect 
on the ordinary assay crucible while the sub-silicates attack it strongly 
to satisfy their desire for silica. 

The student should distinguish between the formation temperature 
of a slag and the melting point of the same slag when already formed. 
It has been shown by Day,’ that when the constituents of a slag are 
finely crushed and intimately mixed as in an assay fusion, the forma- 
tion temperature of the slag is decidedly lower than the melting 
temperature. That is to say, the slag forms without melting and 
actually passes through a pasty stage before coming to perfect fusion. 


Action of Borax in Slags. Borax (Na,B,O; + 10H2O) melts at 
about 560° C. and gives up its water of crystallization forming borax- 
glass. Borax-glass when molten is decidedly viscous and on account 
of its excess of boracic acid it acts as an acid flux. 

! Journal Am. Chem. Soc. 28, p. 1039 (Sept. 1906) 








87 


To determine what relation it bears to silica as regards its acid 
fluxing quality we may consider the matter first from a theoretical 
standpoint, and then from the results of experiments. 

- Considering the borates from their metallurgical classification, 
le. according to the amount of oxygen in the acid to that in the base, 
we may compute the weight of borax-glass necessary to form a mono- 
borate with a unit weight of sodium carbonate and compare this 
with the amount of silica required to form a mono-silicate with the 
same amount of base. From the rational formula for borax-glass 
(Na2O.2B203;) we see that to form the mono-borate (6Na:,0.2B.0s3), 
borax-glass requires five additional molecules of NazO. The equation 
may be written as follows :— 

5NaeCOs + Na,0.2B.03 = 6Na.O0.2B.03 a 5COs. 

Whence we may write the following proportion to find the amount of 
borax-glass necessary to form a mono-borate with 100 grams of soda:— 
5NasCO; : Nae,O.2B.03 = 5a) See 2 =a OU sr, € 
Solving, x is found to equal 38.1. In the same way we may find the 
amount of silica necessary to form a mono-silicate with 100 grams 

sodium carbonate. f 

oes IO, = 212 : 60= 100 : y 
Solving, y is found to equal 28.3. Whence, from the theoretical 
standpoint we may say that in the case of a mono-silicate slag 38.1 
gram of borax-glass is equivalent to 28.3 grams of silica, or when 
borax-glass is used to replace silica in a mono-silicate slag one gram 
has the same effect as 0.743 grams of silica. 

For a bi-silicate slag the relation is different owing to the molecule 
of Na,O already in the borax-glass. The amount of borax-glass 
required to form a bi-borate with 100 grams of sodium carbonate is 
95.3 and the silica for a bi-silica is 56.6. Thus in the case of a bi- 
silicate slag one gram of borax-glass is equivalent to 0.584 grams of 
silica. 

The results of experiments on the size of lead buttons obtained 
in reducing power fusions with varying amounts of silica in some 
instances and borax-glass in others give results approaching the 
theoretical values obtained above. They show that 10 grams of 
borax-glass has the same effect in preventing the reduction of lead 
from litharge as between 6 and 7 grams of silica. 

Rose * in a discussion of the refining of gold bullion with oxygen 
gas made a number of experiments to determine the best proportions 
- of borax, silica and metallic oxides. Borax alone was found to be 


' Lodge Notes on Assaying, 2nd Edit. p. 86. 
2 Inst. Ming. & Metallurgy, 14 p. 396, April, 1905. 


88 


unsatisfactory on account of the rapid corrosion of the crucible. 
Silica alone gave a pasty, very viscous, slag. The best slag found 
corresponded nearly to the formula 3? (Na2O, BeO3) + 38 RO, ¥ 
B:O03, 38102. This is made up according to the following formula, 
9RO + 2NaeB,0; + 98102, where R = Ca, Mg, Pb, u,) Cu, ?Fe, 
3Ni. Leaving out of account the metaborate of soda NagB2QOu, 
it 1s a boro-silicate in which the relation of oxygen in acids to oxygen 
in bases is 2.66 to 1. This slag melts at a low temperature and is 
very fluid at between 1000° and 1100° C. It has only a slight cor- 
rosive action on clay crucibles. The flux contains 3 parts by weight 
of borax glass to 4 parts of silica. 

Charles E. Meyer’ in fluxing zine box slime, made zine into bi- 
silicate with silica and added Na,B,O; for other bases all assumed to 
be Fe.O3; pound for pound, i.e. 1 lb. NagBsO;, for 1 lb. FesQs. 


Fluidity of Slags. It is also necessary to distinguish between 
the melting point and the fluidity of slags. Many slags of low melt- 
ing and formation temperature are entirely unsuited for assay pur- 
poses on account of their viscous nature when melted. As a rule, 
the higher the temperature the more fluid a slag will become, but 
different slags vary much in this respect. All slags are viscous at 
their freezing point, yet one slag will be thinly fluid 200° C. above its 
melting point and another wlll be decidedly viscous at this degree of 


superheat. The viscosity of silicates increases with the percentage 


of silica above that required for the mono-silicate, and the same may 
be said for borates. . 

These silicates and the slags resulting from mixtures of them must 
not be thought of as chemical compounds, but rather as solutions of 
one thing in another, as for instance lead silicate, a solution of silica 
in molten litharge and vice versa. When solidified a slag may be 
either a glass, i.e. a solid solution, or a mixture after the nature of an 
alloy. 


Acid and Basic Slags. Slags more acid than the mono-silicate 
are generally termed acid, while those approaching a sub-silicate are 
called basic. The acid slags are all more or less viscous when molten 
and can be drawn out into long threads. They cool slowly and are 
usually glassy and brittle when cold. The basic slags are usually 
extremely fluid when molten, they pour like water with no tendency 
to string out, in fact they may even be lumpy where the bases are in 
too great excess. They solidify rapidly and usually crystallize to 
some extent during solidification. Basic slags are dull and tough 

1 Jl. Chem. Met. & Ming. Soc. of South Africa, 5 p. 168, Jan. 1905. 





ao.) ee? Le. 


a +. oe a 


se ee 








89 


when cold. They are usually of a dark color and on account of the 
-Orge proportion of bases they contain they usually have a high 
“pecific gravity. 


Mixed Silicates. The mixture of two or more fusible compounds 
usually fuses at a lower temperature than either one taken alone, just 
as for example a mixture of potassium and sodium carbonate fuses 
at a lower temperature than either one of them alone. For this 
reason assayers always provide for the presence of a number of easily 
fusible substances, although their presence is not always necessary 
for the decomposition of the ore. For instance, even in the assay of - 
pure limestone, a base, a certain amount of sodium carbonate also a 
base is always added. 


Use of Fluxes. For the sake of economy in material and time it 
is best to limit the amount of fluxes to the needs of the ore. The 
great saving to be made in this way may be illustrated as follows: 
If we use twice as much flux as necessary, we have to use twice as large 
a crucible, which cuts down the furnace capacity very considerably, 
and besides the large charges require a longer time in the furnace to 
fuse and decompose the ore. 


Slags for Class 1. Siliceous Ores. ‘To fuse a siliceous ore, 
basic fluxes must be added, the alkali carbonates and litharge being 
the ones at our disposal. The bi-silicates of soda and lead being 
readily fusible and sufficiently fluid for our purpose we may limit our 
basic fluxes to the amount necessary to form these silicates. Sodium 
carbonate and litharge combine with silica to form bi-silicates in 
proportions indicated in the following equations :— 

NasCO; a SiO. = Na,SiO; aa COs, 
PbO + S102 = PbSiO, 

From a comparison of the molecular weights of the left hand mem- 
bers of these equations, it may be determined that one assay-ton of 
pure silica will require either 51.5 grams of sodium carbonate, or 108 
grams of litharge to form a bi-silicate. As the mixed silicate of soda 
and lead works better than either one alone it will be wise to make a 
double silicate of the approximate formula Na,O.PbO.28i02. We 
may use 30 grams of soda (3/5) and 45 grams of litharge (2 /5) for 
one assay-ton of silica or any other inversely proportional amounts 
of the two. 

In assaying an ore we must also provide for a lead button to act 
as a collector of the precious metals. A 25 gram button is usually 
sufficient. To allow for this we will need to add 28 grams more of 


90 


litharge (92% lead) and also some reducing agent say 21 grams of 
argols (R. P. 10). 


Our charge will now stand as follows:— 


Ore T Ase. 
Sodium-carbonate 30 Grams 
Litharge for slag 45 gms. | 

Litharge for button 25 gms. \ 70 Grams 
Argols (R. P. 10) 2+ Grams 


So far we have been considering an ideal ore, pure silica. This, 
however, is rarely if ever found in practise. The ordinary so called 
siliceous ore rarely contains more than 90 per cent of silica and often 
not more than 80 or 85 per cent. The rest of the ore will be either 
neutral or basic, usually basic, such as AlO3, Fe.O3, MnOns, CaO, 
MgO, etc. or will be in the form of some sulphide mineral as pyrite, 
which will be converted into a base during the fusion. 

To take care of these bases and still maintain a bi-silicate slag, 
we have the choice of two options; we may calculate how much of 
the silica of the ore will combine with the bases thereof, and then 
supply basic fluxes for the remaining silica, or we may supply basic 
fluxes for all of the silica and then supply acid fluxes, preferably borax 
or borax-glass, to form a bi-silicate or its equivalent bi-borate with 
the bases of the ore. This second method is the one generally fol- 
lowed in practice, that is to say for all siliceous ores a fixed amount of 
soda, borax, litharge and reducing agent is used and then if there is 
much iron, manganese, calcium, magnesium or aluminum oxide pres- 
ent an additional amount of borax is added to the charge, not usually 





or simply to make the slag more acid but more especially to make a ; 


better and more fusible slag with these rather refractory metal oxides. 
The following are examples of bi-silicate charges for siliceous ores, 


Ore 7 ALT. eae 2 Aw. 
Soda (NaeCO3) 15 gms. 30 gms. 60 gms. 
Borax 3-5 5-107 a ee 
Litharge DOC yay Seely rh 
Argols (Ro P: 10)> 22" oe See o 


The larger the amount of ore used the more necessary it becomes 
to keep down the quantity of fluxes. The following charges more 
acid than the bi-silicate are regularly used in this laboratory for the 
assay of siliceous tailings. If the tailings were pure silica the slags 
would be almost tri-silicates. 





91 


Ore eA a AST. DA esl 

Soda (Na2CQOs3) 30 gms. 60 gms. 150 gems. 

Borax ae aeons Lia 

Litharge Bp 00) Meet OY es 

Argols | : for a 25 gram bead button. 


| The results obtained with the last mentioned charges are good, the 
_ slags of course are more viscous than the bi-silicate slags but they pour 
well even when fusions are made in the muffle furnace. The crucibles 
are practically unattacked and can be used for many such fusions 
if of good quality and especially if care is taken to cool them slowly. 

The following table of bi-silicate factors contains all of the data 
— necessary for fluxing siliceous ores with any of the common basic 
fluxes, and it is thought that the above explanation will make its 
use readily understood. 


TABLE XVIII. BI-SILICATE SLAG FACTORS. NO. 1. 





Quantity of Bases Required. 

















SiOz Mie 
PbO | NavCOs | KeCOs NaHCOs 
_ 1 assay-ton 108.4 gms. | 51.5 gms. | 67.1 gms. | 81.7 gms. 
_ 10 grams ele TR eh 23 Oe paiy © 

















One gram of FeO neutralizes 8/10 gms. SiO. or requires 1.4 gms. 
_ borax-glass. One gram of CaCO; neutralizes 0.5 gms. SiO or re- 
- quires 1.0 gms. borax-glass. 

To make a bi-silicate slag and at the same time to keep the quantity 
3 of fluxes at a minimum the procedure would be as follows:—First, 
_ find the amount of silica which will be converted into bi-silicates by 
q the bases of the ore. Second deduce this quantity from the amount 
_ of silica in the ore and add basic fluxes for the remainder. 

For example take an ore of the following composition :—Si0O2 80%, 
— FeO 10%, CaCO; 10%, ore charge 1 assay-ton. 





Silica equivalent FeO 2010 Pia pan ta tater 
o = macs 2.916 x 5 = 1.46 
Total 3.78 

Silica in ore moth 1292166 = 23.33" ema. 

less silica equivalent of FeO and CaCQ; . 3.78 


Silica for which fluxes are to be added 19.55 gms. 






92 . ; ath 


Starting with 30 grams of soda (N 240) ad nog ir grams of 


borax this gives the following charge :— 


Ore LCA 
Sodium-carbonate -. 30- gis: 

Borax 10235 

Litharge for slag 23 BQ 

Litharge for lead button 27 

Argols for 25 gram button. 






Following the more usual custom of adding a given amount of q 


soda-litharge flux and then adding extra borax for the bases, we have 
simply to compute the borax to add for the 10 per cent of ferrous 
oxide and limestone. 


Borax glass equivalent FeO 2.916 X 14 = 4.08 

Borax glass equivalent CaCO; 2.916 xX 1.0 = 2.92 
Total 7.00 

Whence the charge becomes 

Ore DAG. 

Soda (NaeCQs;) 30 gms. 

Borax-glass Oita 1G. 25 

Litharge 602 

Argols for a 25 gram lead button. 


Slags for Glass 1 Basic Ores. In the assay of basic ores we have 
to add acid fluxes, silica and borax to obtain a fusible slag. Also 
on account of the fact that the silicates of iron, manganese, calcium, 
magnesium and aluminum are by themselves infusible, or nearly so, 
at the temperature of the assay furnace, it is always customary to 
add some soda and excess litharge to the charge. These latter, com- 


bining with some of the silica and borax, form readily fusible com- 


pounds which help to take into solution the silicates of the basic 
oxides and by diluting them give more fusible and fluid slags. A 
quantity of soda equal at least to that of the ore is generally taken 
as a starting point, and very often a quantity of litharge equal to 
that of the ore is also allowed for the slag. 

The silicate degree of the slag will depend on the character of the 
bases. For Class 1 ores consisting principally of iron, manganese, 
calcium, magnesium and aluminum it has been found best to approxi- 
mate a sesqui or a bi-silicate slag. 


The following table of bi-silicate slag factors will ee the 
calculation of charges for basic ores. 


a 








by he. 93 
‘ m r ary’ Z : 


oe ap ee “ ; 
- 2 ' TABLE XIX. BI-SILICATE SLAG FACTORS NO. 2. 








“ Quantity of Bases. Quantity of Acid Required. 

1 A. T. FeO 24.3 gms. SiOe 
1 A. T. CaCOs ig at AE . 
1 A. T. MgCOs 21:0 pas “ 
oF 1 A. T. MgO 4574." ss 
10 gms. PbO Din ae ene 
30 “ NaHCOs 10.8 a 4 
30 “ NaeCOs 16.4 - * 
Peet KOO: 4.4 i a 








For sesqui-silicates use three-quarters of the above quantities of 
silica. 

When borax-glass is substituted for silica consider that one gram 
of borax-glass is equivalent to 6/10 gram of silica. The amount of 
silica which should be replaced by borax has not been determined, 
but on account of the greater fusibility and fluidity of boro-silicates 
it is well to replace at least } to $ of the silica with its equivalent 
of borax or borax-glass. 

The following example will illustrate the use of the table. Take 
an ore of the following composition CaCO; 90%, SiOz 10%, ore charge 
1 assay-ton. Starting with 30 grams of sodium carbonate and 50 
grams of litharge, 20 for the slag and 30 for the lead button an jlan- 
ning for a bi-silicate slag the silica requirements of the different bases 
are as follows:— 

The CaCO; of the ore requires 0.9 X 17.6 


15.80 gms. SiO, 


30 grams soda requires 7 Nek: Pee oo 

20 he PbO 2 — 5.4 6 éé 
Total 37.6 

Deducing the silica of the ore 2.9 

Silica or equivalent borax necessary 34.7 


Putting in say 20 grams of silica we have to provide borax equivalent 
to 34.7 — 20 = 14.7 grams of silica or 24 grams borax-glass. 
_ The final charge stands 


Ore LAS TD: Litharge 50 gms. 
Sodium carbonate 30 gms. Argols for 25 gram button 
Borax-glass 24 gms. Silica 20 gms. 


Reducing and Oxidizing. Reducing and oxidizing reactions, 
are common in fire assaying as in other chemical work, and practically 
all fusions are either reducing or oxidizing in nature. For instance, 
the scorification assay is an oxidizing fusion in which atmospheric 


94 


air is the oxidizing agent, while the crucible fusion of a siliceous ore 


is a reducing fusion in which argols, flour or charcoal act as the re- 
ducing agents. 

By reducing power as used in assaying is meant the amount of lead 
that one gram of the ore of substance will reduce when fused with 
an excess of litharge. For instance, if we use 5.00 grams of ore and 
obtain a lead button weighing 16.50 grams the reducing pov of the 
ore is 

16.50 


5.00 
By oxidizing power is meant the amount of lead which one gram 





= 3.30 


of the ore or substance will oxidize in a fusion, or more exactly it is — 


the lead equivalent of a certain amount of reducing agent or ore which 
is capable of being oxidized by one gram of the ore or substance. 

Reducing Reactions. The reduction of lead by charcoal is shown 
by the following reaction :— 


2PbO + C = 2Pb + COs 
From which it is seen that one gram of pure carbon should reduce 
2 X 207 


rahe Bs: 34.5 grams of lead. As however charcoal is never pure 


carbon the results actually obtained in the laboratory will be some- 
what less usually from 25 to 30. All carbonaceous materials have 
more or less reducing power. Those most commonly used as reducing 
agents in assaying are charcoal R. P. = 27.5, argols R. P. 8-12, 
cream of tartar R. P. 5.5, flour R. P. 9 — 12. 

Besides carbonaceous matter many other substances and elements 
are capable of reducing lead from its: oxide. The most important 
of these are metallic iron, sulphur and the metallic sulphides. The 
reduction of lead by iron is shown by the following reaction :— 

PbO + Fe = Pb + FeO 


Whence the reducing power of iron is oe =aed 


The reducing power of sulphur and the metallic sulphides will 
vary, dependent on the amount of alkaline carbonate present. For 
instance, the reduction of lead by sulphur in the absence of alkaline 
carbonates is shown by the following reaction :— 


2PbO +S = 2Pb + SO, 
The reducing power of sulphur under these conditions would be 








95 


In the presence of sufficient alkaline carbonates the sulphur is 
oxidized to sulphur-trioxide which combines with the alkali to form 
a sulphate. The reaction is as follows:— 

From which we see the reducing power of sulphur under these con- 
ditions should be : 

cele 19.4 

32 

In the same way we find that the reducing power of the metallic 
sulphides varies according to the amount of available alkaline car- 
bonate present. For instance, in the absence of alkaline carbonates 
the following equation expresses the reaction between iron pyrite 
and litharge :— 

FeS, + 5PbO = FeO + 5Pb + 280, 
Whence the reducing power is found to be on = 8.6 
- In the presence of an excess of sodium carbonate the sulphur is 
oxidized to trioxide as indicated by the following reaction :— 
FeS, + 7TPbO + 2Na2COs = FeO + Pale + 2NaeSO, of ZC. 


Which gives a reducing power of ae == 2s 


In order to get such a high reducing power as this it is necessary to 
have a very basic slag. 

Any silica present combines with soda and litharge to form a silicate 
and if it is present in any considerable amount the litharge and soda 
are rendered unavailable for the higher oxidation of the sulphur. 

_ The amount of lead reduces from any charge by any reducing agent 
is always a function of the temperature and the silicate degree. Other 
things being equal the more basic the charge the greater the amount 
of lead reduced by a unit quantity of the reducing agent. Thus, 
a certain sample of argols showed a reducing power of 11.04 when 
silica for a sub-silicate was added, 10.93 for a mono-silicate, 10.62 
for a bi-silicate and only 9.26 for a tri-silicate. : 

The accompanying table gives the reducing power of some of the 
common sulphides. The theoretical figures are computed both for 
sulphur oxidized to SO. and SOs. In the last columns are given the 
reducing power of the pure minerals using the following charge NagCO; 
5 gms., PbO 30 gms., SiO. 2 gms., ore to yield an approximate 25 
gram button. 


96 


TABLE XX. REDUCING POWER OF MINERALS. 


























Computed ; 
Mineral Formula || ———--__—__ eee pees d 
S to SO2 | S to SOs 

Cralena PbS 2.6 3.46 3.41 
Chalcocite CuzS 3.9 5.2 

Arsenopyrite FeAsS 5.7 6.96 8.18 
Stibnite SbeSs San 7.30 6.75 
Chaleopyrite CuF es: 6.2 8.44 7.85 
Sphalerite ZnS 6.37 8.5 7.87 
Pyrrhotite Fe7Ss 7.35 9.9 10.00 
Pyrite FeS: 8.6 bo be Or 11.05 





Oxidizing Reactions. Certain metals notably iron, manganese, 
copper, cobalt, arsenic and antimony are capable of existing in two 
states of oxidation. When fused with a reducing agent the higher 
oxides of these metals are reduced to the lower state of oxidation at 
the expense of the reducing agent. Ores containing these higher 
oxides are said to have an oxidizing power on account of this property 
of using up reducing agent. For convenience this oxidizing power 
is measured in terms of lead although the bulk of the oxidizing re- 
action in any assay fusion is probably accomplished against the re- — 
ducing agent of the charge. : 

For instance if in an assay fusion containing silica we have ferric 
oxide, sufficient for a bi-silicate, and carbon the following reaction 
takes place :— 

2Fe.0; + C + 4810, = 4FeSiO; + CO, . 
From which we find that one gram of Fe,O3; requires 0.037 gram of 
carbon to reduce it to FeO. Expressed in terms of lead the relation 
would be as follows :— 
Fe,0; + Pb = 2FeO + PbO 


That is to say the oxidizing power of FegQs is ~ = 13 


Similarly , . 
MnO, + Pb = MnO + PbO 


The oxidizing power of MnO, is = = 2.4, which means that each 


gram of MnO; present in a fusion with litharge and a redueing agent 
will prevent the reduction of 2.4 grams of lead. It is easily seen 
therefore that this oxidizing power of ores must be taken account of 
in computing a furnace charge. The method of determining the 
oxidizing power of ores etc. will be discussed later. 





97 


In the crucible assay of high sulphide ores it is frequently necessary 
to add some oxidizing agent to the charge to prevent the reduction 
of an inconveniently large lead button. A lead button of 25 or 30 
grams is usually sufficiently large to act as a collector of the precious 
metals and were a larger button obtained, it would entail an extra 
loss due to scorification or a prolonged cupellation as well as consum- 
ing extra time in this treatment. When therefore the ore charge 
would of itself reduce more than 25 or 30 grams of lead we ordinarily 
add potassium nitrate (niter) or some other oxidizing agent. Niter 
is almost exclusively used in this country for oxidizing. Its action 
with carbon is shown by the following equation :— 

AKNO; + 5C = 2K20 + 5CO, + 2N2 
From which the theoretical oxidizing power of niter expressed in 
terms of lead is found to be 5.17. The actual oxidizing effect of niter 
is always found to be lower than this due to the acidity of the charge 
and the probably escape of some oxygen. 

In the soda-litharge-silica fusion such as commonly used in actual 
niter assays (sub-silicate), the reaction between niter and iron pyrite 
will be about as indicated by the following equation :— 

k 10OKNO; -f 4FeS. a S10. = FesSiOg oh 5KoSO, + 350, + 5No 
- Assuming the reducing power of pyrite in this type of charge to be 
10.0 the oxidizing power of niter is found to be 4.7. 

A slight oxidizing effect may be obtained by using red lead (Pb3Q,) 
in place of litharge and this is sometimes done especially in England 
and the English colonies. The oxidizing effect of red lead is shown 
by the following reaction :— 

Pb;0, + Pb = 4PbO 


The oxidizing power in terms of lead is wan = 0.30 


Slags for Class 2 Ores. When atisere contains any considerable 
proportion of sulphide minerals and:especially when they are present 
in such proportions that it is necessary-to add niter to prevent the 
reduction of too much lead it will be:¢#eund that the charges recom- 
mended for class 1 ores will not allow..a satisfactory decomposition 
of the ore. Instead of obtaining two:products, slag and lead, as the 
result of the fusion a third intermediate product (matte) is often - 
obtained. This amounts to incomplete decomposition of the ore 
and is a sure indication of low results. The rate of melting down or the 
temperature of the furnace during the early part of the fusion seems 
to have a great. deal to do with the formation of a matte... A rapid 
melting in a hot furnace seems to allow a more complete. oxidation 


98 


of the sulphur, possibly because the litharge is utilized for this pur- 
pose before it is locked up by the silica. At any event as rapid melt- 
ing as may be without boiling over seems to give the most satisfactory 
results with class 2 ores. 

A matte is much less likely to be formed, however, with a less acid 
charge and it has been found best therefore to make a slag approach- 
ing a sub-silicate for all heavy sulphide ores, as by this means more 
uniformly satisfactory results are obtained. : 

The Cover. In practically all crucible assay work it is customary 
to place on top of the mixed charge in the crucible-a cover of some 
fusible substance. Different assayers advocate different materials 
as salt, borax, borax-glass, soda as well as different flux mixtures. 

The idea of the cover is that melting early it makes a thick glaze 
on the sides of the crucible above the ore charge and tends to prevent 
sticking of particles of ore or lead globules which might be projected 
or left there by the boiling of the charge. As the fusion becomes quiet 
and the temperature rises, most of this glaze runs down to join the 
main charge and carries with it any small particles of ore or lead which 
may have stuck to it in the early part of the fusion. 

The salt cover is thinly fluid when melted. It does not enter the 
slag but floats on top of it thus serving to keep out the air and to 
prevent loss by ebullition. 

The borax cover fuses before the rest of the charge. It is thick 
and viscous when melted and serves to prevent loss of fine ore by 
“dusting”, as well as to stop loss by ebullition. It finally enters the 
slag and so ceases to be a cover after the fusion is well under way. 

Some assayers object to the use of salt on the ground that it is 
liable to cause losses of gold and silver by volatilization. It is a well 
known fact that gold chloride is volatile at a comparatively low tem- 
perature, commencing at 180°C. and that silver chloride is volatile 
in connection with the chlerides of arsenic, antimony, copper, iron, 
lead etc. When an ore eentains substances such as manganese 
oxide, basic iron sulphate ete., capable of generating chlorine upon 
heating with salt it would seem wise to omit the use of salt. If it 
is not desired to use salt a good cover may be made from a mixture 
of borax-glass and sodium carbonate in the proportion of 10 Bai 
. of the former to 15 parts of the latter. 

Testing Reagents. Each new lot of litharge and test lead should - 
be assayed for silver and gold so that when any is found to be present 
a proper correction may be made. Different lots of argols, charcoal 
etc. are also found to vary in reducing power so that their reducing 
powers should also be determined. 








99 


The following procedure is designed, Ist, to allow the student to 
determine the reducing power of argols, charcoal or other reducing 
agents and at the same time to determine the silver correction for 
litharge, and, 2nd, to familiarize him with the principal operations 
connected with the crucible method of assay. 

Procedure. Take two E or F pot furnace crucibles, or 12 or 15 
gram muffle crucibles. 

Weigh into them, in the order given the following :— 


INO. 1 No. 2. 
Sodium carbonate 5 grams Sodium carbonate 5 grams 
Silica Dites Silica ae’ 
Litharge iP ge Litharge OUe 
Argols Qeee eye: Me harcoal Lat 


Weigh the argols and charcoal on the pulp balance as exactly as 
possible, the others on the flux balance. Mix thoroughly with the 
spatula by turning the crucible slowly with one hand while using 
the spatula with the other. When finished tap the crucible several 
times with the handle of the spatula to settle the charge and to shake 
down any material which had lodged along the inside of the crucible 
above the charge. Finally put on a 4 inch cover of salt. 

Pot Furnace Fusion. Have a good bright fire in the pot furnace 
which should not however be filled with coke more than half wuy to the 
bottom of the flue. Figure to so place the crucibles that their tops 
shall not be much above the bottom of the flue. Place a piece of 
cold coke directly under each crucible as it is put into the furnace. 
Cover the crucibles and pack coke around them being careful to pre- 
vent the introduction of any coke or dust. Close the top of the fur- 
nace, open the draft if necessary and urge the fire until the charges 
begin to fuse. Then close the draft and continue the melting slowly 
enough to prevent the charges from boiling over. When the charges 
have finished boiling, note the time and open the draft if necessary 
to get a yellow heat and continue heating for 10 or 15 minutes. 

Pour the fusions into the crucible mold, which has been previously — 
coated with ruddle, thoroughly dried and warmed. When cold, a 
matter of 5 or 10 minutes for a small fusion, break the cone of lead 
from the slag and hammer it into a cube to thoroughly remove the 
slag. Weigh the buttons on the pulp balance to the nearest tenth 
of a gram and record the weights and reducing powers in the note 
book. — 

Save the lead buttons and cupel them, using. eee cupels. They 
should contain all of the gold and silver in the 60 grams of litharge 
used. Weigh the beads and part to see if gold is present. Record 


_— 


VS 


100 


the weights of the beads and compute the correction for silver in 30 
erams of litharge. 

Muffle Fusion. If the fusions are to be made in the muffle have 
the muffle red and the fire under such control that the muffle can be 
brought to a full yellow in the course of a half hour. Melt at suffi- 
ciently low temperature to avoid violent boiling and then raise the 
temperature and pour as in the case of the pot furnace fusion. 

Notes. 1. So-called silver free litharge can now be purchased but even this 
often carries traces of gold and silver. 

2. In assaying samples of litharge low in silver 120 to 240 grams may be required 
to give a button of sufficient size to handle and weigh. 


3. It is convenient to use litharge in multiplies of 30 grams and therefore the 
silver correction is based on 30 grams of litharge. 


4. The temperature which the muffle should have before the crucibles are in-’ 


troduced depends upon the number of charges which are to be put in at one time. 
If only one or two the temperature should be low to avoid danger of boiling over. 
If, however the muffle is to be filled with crucibles the initial temperature may be 
higher as the crucibles can be depended upon to decidedly lower the temperature. 

5. Pour the fusions carefully into the center of the molds and do not disturb 
until the lead has had time to solidify. 


The following are the reducing powers of some of the common re- 
ducing agents. 


Charcoal 23-30 Corn Starch 11.5-13 
Argols 7-12.5 White Sugar 14.5 


Flour 12-15 Cream of Tartar 4.5-6.5 


Assay of Class 1 Ores. Gold or Silver. 


This is the most common class of ores and as it is also the one which 
presents the fewest difficulties for the assayer, it is considered first. 
Actually, ores with no traces of sulphides are somewhat of a rarity 
but the methods given below may be adapted to ores containing mod- 
erate amounts of sulphides by simply decreasing the amount of re- 
ducing agent used. 

Procedure. Carefully van some of the ore, estimate and record in 
the note-book the amount and character of each of the slag forming 
constituents and also of any sulphides present. If the ore is mainly 
siliceous weigh out one of each of the following charges :— 


Charge (a). Charge (b). 

_) Ore 0.5 A. T. Ore 0:5 AT 
Sodium carbonate 30 grams Sodium carbonate 15 grams 
Borax eo) ne Borax 3-5 
Litharge SOR Litharge BQ: as 
Argols ‘i ~  Argols - 


* Argols enough combined with the reducing material of the ore to 
give a 25 gram button. 





101 


_ Weigh out the fluxes and place in the crucible in the order given and 
finally add the ore and argols last of all. Weigh the argols and ore on 
the pulp balance, the others on the flux balance. Mix thoroughly 
and place a 3 inch cover of salt or soda-borax mixture on top. 

Use F pot furnace crucibles, 15 or 20 gram muffle crucibles, if 
work is to be done in the muffle. : 

Fuse at a moderate red heat to avoid danger of the charge boiling over 
and when quiet raise the heat to a bright yellow. Allow 15 minutes 
of quiet fusion. Pour as usual, tapping the crucible gently against 
the mold if necessary to insure getting out the last globules of lead. 

When cold, separate the lead buttons from the slag keeping them 
in order (a) (b). Record in the note-book the character and appear- 
ance of the slags, the ease or difficulty of the separation of each 
from the lead buttons, the appearance of the lead buttons and their 
greater or less malleability. 

Weigh the lead buttons on the flux balance and cupel carefully 
to obtain feather litharge. Weigh the silver beads, correct for silver 
in the lithage used, part and weigh any gold found and finally report 
the value of the ore in oz. per ton. 

Both of these charges should give good results on a siliceous ore. 

Charge (a) is a little less expensive, but charge (b) is more commonly 
used, as the slag contains two bases and the excess litharge will hold 
a moderate amount of impurities in solution. Charge (b) also gives 
a better separation of lead button and slag and has the further ad- 
vantage that if the ore contains slightly more sulphides than was 
estimated the litharge will take care of them, giving a lead button 
free from matte. In charge (a), if we have more carbonaceous re- 
ducing agent plus sulphide mineral than the 30 grams of litharge can 
oxidize some of the sulphur will combine with various metals of the 
charge, principally lead, and form a matte which will appear immedi- 
ately above the lead button. 
_ Approximately 28 grams of litharge from each charge will be re- 
duced to give the 25 gram lead button and is therefore not available 
to combine with the silica. The active! fluxes are then in charge 
(a), 30 grams of soda, 3-5 of borax, 2 grams of litharge and a little 
-K.O from the argols, totaling approximately 23 times the ore. In 
charge (b), the active fluxes are 15 grams of soda, 3-5 of borax, 22 
grams of litharge and a small amount of K.O, totaling approximately 
3 times the ore. A very good rule to follow in making crucible charges 
is always to use at least 25 times as much active flux as ore. 


! By active fluxes is meant a flux which is to appear in the slag and therefore 
does not include the litharge which goes to form the lead button. 


102 


Borax in the charge should be increased as the bases increase. For 
an ore with 10 or 20 per cent of iron, manganese oxide or limestone 
add up to 10 or 15 grams of borax or 5 to 8 grams of borax-glass. 


Notes. 1. Some assayers prefer to omit the borax from the charge and use a 
cover of crude borax or borax-glass in place of the salt. A borax cover may be used 
to advantage with ores which “dust” in the crucible, as the borax swells and melts 
early tending to catch and hold down the fine particles of ore which are projected 
upward from the charge. 


2. The crucible should never be more than two-thirds full when the charge is 
all in. 


3. If a silver assay alone is asked for it is customary to omit parting and report 
the combined precious metals as silver. 


4. In assaying for gold alone if sufficient silver for parting is not known to be 
present, a piece of C. P. silver should always be added to the crucible or to the lead 
button before cupelling. If the approximate amount of gold is known allow about 
eight times its weight of silver. 


5. The slag should be fluid on pouring and should be free from lead shot. If 
it strings out in long threads on pouring it is too acid. If it is pasty or lumpy, 
either the fusion has not been long enough to thoroughly decompose the ore, or the 
charge is too basic and more borax and silica should be added. ‘The crucible should 
have a thin glaze of slag and should be but little corroded. It should show no 
particles of undecomposed ore or ‘‘shots’ of lead. These latter can best be seen 
immediately after pouring and the student should make it a point to examine his 
crucible immediately after every pour. Neither the cover nor the outside of the 
crucible should show any glazing, as this indicates that the fusion has boiled over. 
The cold slag should be homogeneous, as otherwise it indicates incomplete decom- 
position of the ore. Glassy slags are usually preferred by assayers but are not 
essential to all ores. 


6. If the button is hard or brittle or weighs more than 30 grams it should be 
scorified before cupelling. Hard buttons indicate the presence of copper, antimony, 
or nickel. Brittle buttons may be due to antimony, arsenic, zinc, sulphur, litharge 
or may be rich alloys of lead and the precious metals. 


7. Examine carefully the line of separation of the slag and lead. The separation 
should be clean with no films of lead adhering to the slag. There should be no third 
substance between the slag and lead, nor should the surface of the lead show any 
disposition to crumble when hammered. Any lead gray, brittle substance between 
the lead and slag or attached to the lead button is probably matte. This indicates 
incomplete decomposition of the ore due to too short a time of fusion or to incorrect 
fluxing. If from the latter cause, decreasing the silica and increasing the soda and 
litharge will usually prevent its formation in a subsequent fusion. 


8. For low grade gold ores or tailings a 1 A. T. charge is commonly used and for 
very low grade ores and tailings from 2 to 10 A. T. are run usually in the pot furnace. 
In using large charges of siliceous tailings the following charges more acid than the 
bi-silicate have been used with excellent results. 


Gold Ore | Low Grade Tailings 


Ore LAE 2A. T. 5 Age 
Sodium carbonate 30 grams 60 grams 150 grams 
Borax 3-5. CS 6-10. “ 15-25 “ 
Litharge HO ties 9005 180 55 
Argols for a 25 or 30 gram button in each case. 

Crucible G, 20 or 30 gm_ “4H, 30 or 35 gm K. 








103 


Assay of Class 2 Ores 


Ores of this class containing only small amounts of sulphides are 
assayed exactly the same as class 1 ores using lesser amounts of argols. 
When however, sulphides are present in such amounts as to reduce 
a lead button too large to cupel (i.e. over 25 or 30 grams) a different 
method of procedure must be followed. The most important methods 
for the assay of these ores follow:— 


1. Scorification. This method has. already been considered. 
It is not well suited for gold ores and fails for many silver ores. 


2. Niter Method. The reducing power of the ore is first de- 
_termined by means of a preliminary assay. Using the figure thus 
obtained a certain amount of niter is added to the regular fusion to 
oxidize a part of the sulphur of the ore, thus preventing the reduction 
of too large a lead button. This is perhaps the most common method 
for the assay of sulphide ‘ores. The sulphides are decomposed partly 
by litharge and partly by the niter. 


3. Iron Method. The litharge added to the charge is kept low 
so that the lead from it plus that in the ore will yield a button of 
suitable size for cupelling. The sulphide minerals of the ore are 
decomposed by the means of the metallic iron. This is a very good 
method for many ores and is very commonly used. 


4. Roasting Method. A carefully weighed portion of the ore 
is roasted to eliminate sulphur, arsenic, antimony etc. and the roasted 
ore is then assayed as a class 1 ore. 


5. Combination Wet and Fire Method. The sulphides ete. 
of the ore are oxidized with nitric acid, the silver is precipitated as 
chloride and combined with the insoluble residue containing the gold, 
is assayed by either scorification or srucible. 


Niter Assay. Preliminary Fusion. Procedure: Take from 
21 to 10 grams of ore depending on ‘the amount of sulphide present, 
2} grams for pure pyrite, and correspondingly greater amounts for 
ores containing less sulphides. If the ore is mostly galena as much 
as 7 grams may be taken, the idea being always to get a button of 
25 or 30 grams. (See reducing power of minerals.) Take the same 
amount of sodium carbonate as ore, 60 grams of litharge and up to 
5 grams of silica. If the ore contains silica a proportionately less 
amount should be added. Use an E crucible for the pot furnace or 
a 12 or 15 gram crucible for the muffle. Weigh out the fluxes first, 


104 


in the order given and place the ore on top mixing thoroughly with a 
spatula. Place a 4 inch cover of salt on top. 


Fuse for 10 or 15 minutes finishing at a good yellow heat. Pour — 
into crucible mold, allow to cool, separate the lead from the slag and 


weigh on the pulp balance to tenths of grams. Divide the weight of 
the lead by the weight of the ore taken to obtain the reducing power. 

It should be noted that this reducing power is not an absolute 
thing but depends upon many factors such as the ratio of sodium 
carbonate to ore, the amount of borax, litharge and silica added as 
well as the temperature at which the fusions are conducted. The 
size of the lead button reduced in any fusion is decreased by any in- 
crease of the borax and silica and is increased by any increase of the 
litharge, soda or temperature. 

The charge suggested for determining the reducing power of an 
ore gives as a rule slightly higher results than are obtained in the regu- 
lar fusion, due to the large amount of litharge used in the preliminary 
fusion. ‘This seems to be necessary. however to insure the presence 
of a sufficient excess of litharge for all ores. The reducing power 
obtained in the regular assay is called the working reducing power, 
to distinguish it from that obtained in the preliminary fusion. 


Estimating the Reducing Power of Ores. In many instances 
it is possible to estimate the reducing power of an ore within close 
limits. This requires a knowledge of the reducing powers of the com- 
mon sulphide minerals (see Table XX), as well as the knack of 
vanning. The ore is vanned and the per cent of the various sulphides 
estimated, from which data the reducing power is found. For in- 
stance, if the ore is 50% pyrite and the rest gangue, the reducing power 
will be about 5.5 (50% of R.P.wf pure pyrite). If it is 40% galena 
and 10% sphalerite, the redueimggpower will be 40% of 3.4 + 10% of 
7.9 = 2.15 approximately. "The reducing power of the ore being 
equal to the sum of the products of the reducing powers of the differ- 
ent constituents, multiplied’ by? tHe percentage of each | in the ore and 
the whole divided by 100.  “#* # . 

In general if the amount of Stlphides in the ore is comparatively 
small and especially if only 0.5 ‘assay-ton of ore is used, it is a very 
simple matter to’ obtain a lead’ button of suitable size for cupelling 
by thismeans. If for example we have a mixture of galena and gangue 


mineral containing 50% of pews the reducing power of the ore will be 


ely = 4. 70. aside + assay ae of this ore we heala ebtnin & lead 


9 : ; ; : i 
make Peep: tt as) 3; ‘WES Waa Fog 





a oe 


Se a 


| 
| 











105 


button weighing 24.8 grams without either argols or niter. If we 
had estimated the galena at 40%, we would have added 4 gram of 
argols (R. P. 10) and would have obtained a 29.8 gram button which 
could still be cupelled. In a similar manner if we had estimated the 
galena at 60%, we would have added about | gram of niter and would 
obtain a button of about 20.8 grams, which is also all right for cupella- 
tion. 

When we have practically pure sulphides, as in the case of pyrite . 
or galena concentrates it is again easy to estimate the reducing power 
and properly control the size of the lead button. 


Determining the Oxidizing Power of Niter. The oxidizing 
power of niter is found by fusing & weighed amount with an ore whose 
reducing power is known. To obtain comparative results the slags 
must be exactly like those used for the reducing power fusion and 
moreover to obtain the proper size of lead buttons in the final assay 
the slag that is made there must be similar as regards acidity, litharge 
excess etc. to that made in the preliminary fusion. 

The following example illustrates the method of finding the oxidiz- 
ing power of niter:— 





Ore 5 grams 5 grams 
Sodium carbonate ee Dears 
Litharge 60“ hee 
Niter ess 
Silica Spe tas See 
Lead obtained 24.31 grams _— 6.61 grams 
2 2A-3 1 bt / a 
Reducing power of ore = 4.86 


Lead oxidized by 4 grams of niter 24.31 — 6.61 = 17.70 


he ' 1s 
Oxidizing power of niter = ieee 4.42 


if} ng 
09 


Niter Assay. Regular Fusion. Knowing the reducing power 
of the ore and the oxidizing power of niter we are ready to make up 
the charge for the regular assay. As in the case of class 1 ores it 
seems best always to use at least as much normal sodium carbonate 
as ore and we may start off on this basis. More litharge is used in 
this assay than in those previously discussed and assayers usually 
increase the litharge in proportion to the sulphur. The rule for the 
use of litharge proposed by Lodge is a good one and calls for 20 per 


106 


cent in excess of the amount required to satisfy the reducing power 
of the ore. On account of the large amount of sulphur present a 
matte is often obtained if the acidity of the charge is not carefully 
controlled. It is therefore best in adding silica and borax to avoid 
using more than required for a sub-silicate. 

The following examples may serve to show the method of computing 
charges. They are all based on 0.5 assay ton of ore as that is usually 
the maximum amount used for the niter assay. The litharge is 
computed according to Lodge’s rule, then the silica required for the 
ore, soda, niter and active litharge is found and the rest of the com- 
putation is exactly like that discussed under class 1 basic ores. 


No. 1. No. 2. No. 3. 
90% Galena 50% FeAsS 90% Fes: 
10% CaCO; 50% SiO» 10% SiOz 
Rebs 15 R. P. 4.10 R. P. 9.50 
Ore C.5eA TL: Ob Ass 0.5°A, Te 
Sodium carbonate 15 grams 15 grams 15 grams 
Borax-glass Dat an ae 10s ee i 
Litharge 1 Ue ae $0 es 180-23 
Niter (O. P. 4.2) Sarees: Siar 26.5: * 
Silica = = 5.025 


Procedure: Make up charges for your ores according to the rules 
outlined above. Conduct the fusions as for class 1 ores taking par- 
ticular care when much niter is used until the boiling period is passed. 
As soon as all danger of boiling over has passed, heat rapidly to a full — 
yellow and pour after 20 minutes of quiet fusion. a 

The following table of sub-silicate slag factors will aid in ae de- 7 
termination of the quantity of acids to add. . 


TABLE XXL. SUB-SILICATE SLAG FACTORS. 





Quantity of Acids Required 





Quantity of. Bases 


Silica - Borax-glass 
LANE. | FeO. . oy: 6.08 gms. 7.4 gms. 
LAST. CaCOs 440 oe aac 
1A.T. MgCOs OALO (AOS 
10 gms. PbO 0.67 S -0:82. 
30 . NaeCOs 41 ret Oar 
40 “ NaHCOs ee ree 4a 
10. “2 7 KsGOs 1 125s Ll. 








107 




























Notes. 1. It has been found possible to use somewhat less litharge than that 
necessary to satisfy the reducing power of the ore when a greater proportion of alkal- 
ine carbonate flux is used. This is probably due both to ‘the oxidizing effect of COs 
at a high temperature and to the solvent power of basic alkaline slags for matte. 

2. As sulphide ores usually contain more or less copper, nickel, arsenic, antimony, 
zinc, tellurium and other so-called impurities, the large amount of litharge used 
serves the double purpose of helping to decompose the ore by oxidizing the sulphur 
and associated metals, and also tends to prevent these metal impurities from enter- 
ing the lead button. 

3. Aside from the inconvenience of the preliminary fusion the principal objection 
to the niter assay is due to the low results yielded. This is undoubtedly caused by 
the oxidation and slagging of the precious metals, particularly silver. To avoid 
this source of error only small amounts of niter should be used. When silver alone 
is being sought the niter may be entirely done away with by reducing the ore charge 
to a quantity sufficient to give a lead button weighing between 20 and 30 grams. 
In gold assays, however, a charge less than 0.5 assay ton is to be avoided as it fails 
to give a sufficiently close valuation of the ore. 

4. Part of the oxidized precious metals may be recovered from the slag after the 
fusion is quiet by the addition of some reducing agent. For instance, if the fusion 
has been made in the muffle and without salt covers some crucibles of soft coal may 
be placed ni the mouth of the muffle after the fusions have become quiet. The 
smoke filling the muffle will enter the crucibles and reduce some lead from the slag 
which will in turn take with it part of the silver and gold. 


The Iron Assay. The iron nail thethod of assaying sulphide ores 
is radically different from any of the other methods described. The 
principal difference being that metallic iron, usually in the form of 
nails, is used as the reducing and desulphurizing agent. As iron 
reduces lead from litharge this latter reagent is limited to 30 grams 
or less and to make up for this the quantity of alkali carbonate is 
increasing to two or three times that of the ore. Just before pouring 
the excess iron is removed. | 
~The chemical reactions taking place in the crucible are entirely 
different from those of the other crucible methods. In the case of 
-the argols, niter and roasting methods of assaying, the sulphides 
of the ore are oxidized by litharge, niter or the oxygen of the air and 
the sulphur either passes off as SO2 and SQ; or in the presence of sodium 
carbonate is converted into sodium sulphate which floats on top of 
the slag. In the iron assay, part of the sulphur is oxidized by the small 
amount of litharge used and the rest stays as sulphide, appearing 
either as an iron matte on top of the lead button or dissolved in the 
excess of basic slag. | 


The following reactions are illustrative of the chemical changes 
which take place. They are arranged i in order of their occurrence.— 


feo | PHS'-+-i 2PbO'= 3Pb + SO. 
Bee CuS + 2PbO°= 2CuPb + SO. 3 
FeS, + 5PbO = 5Pb + FeO + 2580, 
Fe + PbO = Pb + FeO 





108 


when the litharge is all reduced the following occur 


PbS + Fe = Pb + FeS (matte) 
FeS. + Fe = 2FeS (matte) 
SboS3 + 3Fe = 28b + 3FeS (matte) 
AsoS3 + 13Fe = 2Fe;As (speiss) + 3FeS (matte) 
Cus + Fe = Cue + FeS (Partial) 


Finally, if there is a sufficient excess of alkali flux used, the iron 
matte is dissolved by this basic slag, probably as a double sulphide 
of iron and sodium or potassium. 

From the equations it will be seen that copper, arsenic and antimony 





are reduced, at least in part, and go into the lead button, orin the case ~ 


of arsenic form a speiss which ordinarily carries some of the precious 
metals. In general it may be said that the process is not suited for 
ores carrying much nickel, copper, cobalt, arsenic, antimony or 
tellurium. One or two per cent of copper in an ore does not interfere 
seriously with the assay, but when much more than this is present 
some other method should be chosen. Ores containing nickel are 
least of all suited to the process. 

The slag made should not be more acid than a mono-silicate and 
probably a sub-silicate is better for most ores. Occasionally a matte 
is found-on top of the lead button and this generally contains more or 
less gold and silver. It indicates a too acid slag, an insufficient 
amount of alkaline flux or too short a fusion. The slag although 
basic does not attack the crucible to any extent and crucibles may 
ordinarily be used a number of times. 

The method is a most excellent one on suitable ores and the author’s 
experience has been that in nine cases out of ten, students will obtain 
considerably higher results using this method than the niter method. 
It has the advantage that no preliminary fusion to determine the 
reducing power is necessary and that if the lead of the ore is allowed 
for, a button of the proper size for cupellation may always be obtained. 
As before mentioned the method is limited to pure ores and occasion- 
ally a hard button, or speiss may be obtained when no copper, antimony 
or arsenic was suspected in the ore. . Occasionally also, but only when 
the slag is not properly constituted, or when the temperature of fusion 
is too high there may be difficulty in separating the lead from the slag 
and sometimes a thin film of lead may adhere to the slag when the 
two are broken apart. The only other objection is the difficulty 
of removing the nails free from shots of lead, but in general when the 
fusion has gone far enough this will not cause serious inconvenience. 

Procedure: Pan the ore, estimate and record its mineral composi- 


109 


tion. Note especially the per cent of lead minerals. Use a G pot 
furnace or a 20 or 30 gram muffle crucible and weigh out one of the 
following charges. 


+ Galena 
Pure Galena 3 Pyrite Pyrite 
Ore Go.AtrL. HSS ple ei Ge O.55A, ol 
Sodium carbonate 30 grams 40 grams 50 grams 
Borax A eid in. 20 Oe a 
Litharge 79 he aa ets 
Silica Aaa Aa Sas 


Nails;from 3 to 5 (twenty-penny) cut nails or preferably one 3” 
to 4’’ track spike inserted point downward. 
Cover Salt or borax-soda mixture. 

Heat gradually to fusion, fuse from 40 to 60 minutes. Examine 
the nails occasionally and if badly eaten add several fresh ones, leav- 
ing the old ones in the crucible if they cannot be removed free from 
lead. Fuse until the nails may be freed from lead by tapping them 
gently and washing them around in the slag. Remove all nails and 
pour as usual. The slag will be black and should separate easily 
from the lead button. 


Notes. 1. If the ore contains two or more grams of silica none need be added. 
2. If bicarbonate of soda is substituted for the normal carbonate use a currespond- 


a ingly greater weight. 





3. This fusion requires a somewhat longer time than the niter fusion owing to 
the fact that time must be allowed for all of the charge to come in contact with the 
surface of the iron nails. 


4. The lead may not start to drive in cupelling quite as rapidly as other buttons 
owing to a small amount of iron which is often present. 

The Roasting Method. This method of assaying sulphide ores 
is rarely used, but might be used to advantage on very low grade 
pyritic ores, so will be briefly described. 
| Procedure: Take from 0.5 to 5.0 assay tons of ore and spread out 

‘in a well chalked roasting dish of sufficient size to allow of stirring 
without loss. Have the muffle at a dull red only and the fire so low 
that the temperature of the muffle may be held stationary or raised 
but slowly. Place the dish in the muffle and cover it if the ore con- 
- tains minerals which decrepitate and keep it covered until danger 
from this source is passed. The ore should soon begin to roast. 
When fumes are noticed coming from the ore, check the fire and hold 
it at this temperature for some time stirring frequently. After all 
danger of fusing is over gradually raise the temperature stirring at 
intervals of 20 minutes or one-half hour. Finally heat to about 700° 
GC. for one-half hour, when if the ore contains only sulphides of iron 
and copper, practically all of the sulphur will be removed. If there 


110 


is any doubt about the roast being complete, remove from the muffle, 
add a small amount of charcoal and see if any odor of sulphur dioxide 
is noticed. If the ore contains zinc, a much higher temperature will 
be required to break up the zine sulphate. It is not best, however, 
to carry the roasting temperature above 700° C. 

If the ore is principally galena or stibnite, add an equal weight of 
fine sand or assay silica before commencing the roasting, which should 
be done at a very low temperature to prevent the fusion of the sul- 
phides. 

If the ore contains arsenic or antimony, the roasting operation is 
more difficult. The best conditions for the elimination of these ele- 
ments are alternate oxidation and reduction at a low temperature. 
The presence of sulphur aids in the elimination of these elements due 
to the fact that their sulphides are volatile. To obtain the reducing 
action necessary for the elimination of arsenic and antimony take the 
partially roasted from the muffle, allow it to cool for a few moments, 
and then mix powdered charcoal or coal dust with it and roast at a 
dull-red heat until the coal is burned off. Then add more coal and 
re-roast. Repeat this until no more fumes of arsenic or antimony 
are noticed, then heat with frequent stirring to about 700° C. 

After the ore is roasted, the dish is carefully cleaned out and the | 
ore is charged into a crucible with fluxes and treated exactly as a 
class 1 ore. If the sulphide mineral was mostly iron, the ore will 
probably be found to have a slight oxidizing power due to the forma- 
tion of FeO; and Fe;Q, in the roasting. 

The roasting method of assaying is slow and takes up much muffle 
space. It is open to the liability of serious mechanical and vola- 
tilization losses. Its most useful field would seem to be the assay 
of low-grade pyritic gold ores where a very accurate determination 
of gold is desired. The method usually gives low results in silver. 3 

The combination wet and fire assay is used principally for the de- 
termination of gold and silver in copper and nickel matte, copper 
bullion, ete. A description of the method will be found in the chapter 
on bullion assay. 


Assay of Class 3 Ores. 


The principal ores belonging to this class are those containing some © 
of the higher oxides of iron or manganese, i. e. Fe,O3, Fe3O1, MnQOsz. 
These are reduced by carbon and tend to enter the slag as ferrous and 
manganous silicates respectively. If a charge was made up for these 
ores using only the ordinary amount of argols this might be all used 





111 














up in reducing the oxides of the ore and no lead button would result. 
To remedy this the oxidizing power of the ore should be known be- 
fore making up the charge. 

To determine the oxidizing power of an ore, fuse a known weight 
of it, say 10 or 20 grams with a regular crucible charge for that amount 
of ore and a carefully weighed amount of argols of known reducing 
power sufficient to more than oxidize the ore. The weight of lead 
then, that the argols could reduce from an excess of litharge, minus 
the weight of lead obtained is evidently the amount oxidized by the 
ore. This weight divided by the weight of ore taken gives the oxidiz- 
ing power. 

Having determined the oxidizing power of the ore the assay is 
made in the same manner as for class 1 ores with the addition of the 
extra argols required. 

The following table shows the proper size of crucibles for different 
charges. 


TABLE XXII. SIZE OF CRUCIBLES FOR VARIOUS CHARGES. 

















SSS a 


ae ; ' Furnace 
ES aa a Character of Ore 
Ore Taken Pot Muffle 
Age asl Siliceous ia 15 or 20 gms. 
ma Ad 4 G 20 or 30 “ 
SAT iM H 30 or 35“ 
eG A. TL i Ix 
ee AST Basic. Iron or Niter Fusions | G 20 or 30 “ 
2 1 A. T (4 res ‘ (<3 ce | H 30 66 


CHAPTER IX. 
SPECIAL METHODS OF ASSAY. 


The Assay of Telluride Ores. The determination of the precious 
metals in ores containing tellurium has always been considered more 
than ordinarily difficult. Results obtained by different assayers 
and even duplicate assays by the same man were often widely diverg- 
ent. The literature of telluride ore assaying is extensive and none 
too satisfactory; however, it is safe to say that most of the reported 
differences between duplicates and different assayers have been due 
more to difficulties in sampling than to the chemical interference of 
the element tellurium. When it is considered that most of the tellur- 
ide ores which are mined contain less than 0.1 per cent telluride 


mineral, it is apparent that more than ordinary care must be taken - 


to insure obtaining a fair proportion of this in the final assay portion. 
The telluride mineral itself may contain 40 per cent of gold, so that 
one 100 mesh particle more or less in the assay portion may make a 
difference of several hundredths ounces of gold to the ton. To ob- 
viate as far as possible this lack of homogeneity, all telluride ores 
should be pulverized to at least 150 and preferably 200 mesh and 
then very thoroughly mixed before the assay portions are weighed 
out. | 
Effect of Tellurium. Tellurium is a close associate of both gold and 
silver and is difficult to separate from these metals either in the 
crucible, scorification or cupellation processes. It is not however 
often found in abundance, and even in high grade ores tellurium it- 
self is found in comparatively small amounts. For instance, in two 
high grade ores used by Hillebrand and Allen in their experiments on 
the assay of telluride ores, containing respectively 15 and 19 oz. of 
gold per ton, there was tellurium amounting to 0.074 and 0.092 per 
cent respectively. It seems unreasonable to expect such small quan- 
tities of any element to influence seriously the results of a fire assay. 
In order to study the effects of tellurium in the gold and silver assay 
it is necessary to experiment with ores or alloys containing much more 
tellurium than those above mentioned. The following facts regarding 
the behavior of tellurium in cupellation and fusion are mostly due to 








113 


the work of Holloway,' Pease’ and Smith, whom we have to thank 
for co-ordinating and elucidating much information which was hitherto 
much scattered and of doubtful value. 

Hffect of Tellurium on Cupellation. The presence of tellurium in 
a lead button causes a weakening of the surface tension of the molten 
metal. The result is the metal tends to “‘wet’’ the surface of the cupel 
and this allows particles of alloy to pass into the cupel and others to 
be left behind to cupel by themselves on its surface forming minute 
beads. In the case of a button containing 10 or more per cent of 
tellurium with an equal weight of gold or silver, complete absorption 
may take place. As the proportion of lead in the alloy is increased, 
the amount of absorption becomes less, until when the lead: amounts 
to 80 times the tellurium very little loss of precious metal occurs in 
a properly conducted cupellation. (Smith). 

Tellurium is removed comparatively slowly during cupellation 
particularly in the early stages, as might be expected on comparing 
the heat of formation of its oxide with that of lead oxide. Roses 
gives the following figures for the heat of combination of these metals 
with 16 grams of oxygen,—Pb to PbO 503 Cal., Te to TeOz 386 Cal. 
_ To avoid danger of undue loss in cupellation of buttons from the assay 
of such ores, as much as possible of the tellurium should be removed 
prior to cupellation. It is also evident that large lead buttons (30 
or more grams) should be allowed for in order that the ratio of lead 
to tellurium be high. 

Silver in the alloy protects gold from losses due to the presence of 
tellurium. It appears to act as a dilutant for the gold and should 
always be added to every gold assay for this reason if no other. 

In the case of imperfect cupellation, tellurium is retained by the bead 
and gives it a frosted appearance. In perfect cupellation the final 
eondition of the tellurium is that of complete oxidation to TeOs. 
Owing to its effect in reducing surface tension, resulting often in minute 
beads being left behind, it would be well to use a cupel having a finer 
surface when cupelling buttons containing tellurium. Smith states 
that the loss due to sub-division and absorption in this case is much 
less when a “patent” (magnesia) cupel is used. Losses of gold and 
silver by volatilization during properly conducted cupellation of lead 
buttons from ordinary telluride ores is extremely small. 

Effect of Tellurium in Fusions. Tellurium was formerly thought to 
1 The assay of Telluride Ores. G. T. ESTEE and L. E. B. Pease, Trans. 
I. M. M., 17 p. 175. 

2 The Behavior of Tellurium in Assaying, Bynes W. Smith, Trans. I. M. M., 


17 p. 463. 
3 Trans. Inst. Min. & Met. alt p. 384. 


114 


be oxidized to the di-oxide during fusion and to go into the slag as 
a sodium or lead tellurate. Smith disagrees with this and argues 
that tellurates are decomposed at a red heat, and that lead tellurate 
is white, while he found the litharge slags obtained in the fusion of 
telluride compounds to be black. He believes that tellurium exists 
in the slag as the black monoxide (TeO). 

The slag best suited to the oxidation and retention of tellurium 
in crucible assaying is a basic one containing a considerable excess 
of litharge. The temperature of fusion should be moderately low 
as a high temperature prevents the satisfactory oxidation and slagging 
of the tellurium, probably owing to the formation of lead silicates 
before the litharge has had time to oxidize the tellurium. Smith 
gives the following reaction for the oxidation of tellurtum :— 

2 PbO + Te = PhO + TeO 7 

In support of this he claims to have found the black sub-oxide of 
lead in the slag. 

Practically all authorities agree that the scorification process is 
not reliable for telluride ores. When a button from a crucible assay 
contains too much tellurium for direct cupellation Smith recommends 
fusing or ‘soaking’ the button under an ample amount of litharge 
at a moderate temperature (700—900° C.). 

Hillebrand and Allen used the following charges for ores containing 
from 15 to 19 oz. gold and 0.074 to 0.092 per cent tellurium. 

Ore TPAC Litharge 1380 grams 
Sodium carbonate 30 grams Reducing agent for 25 gram buttons 
Borax-glass The Silver 24 to 3 times gold. 

They find the slag losses no higher than with ordinary gold ores 
and no serious cupellation losses. With ores containing much more 
tellurium than the above, the quantity taken should be reduced and 
the rest of the charge maintained as before. 


The Assay of Ores and Products High in Copper. Crucible 
methods for the assay of matte and ores high in copper have largely 
supplanted the older scorification method. This is due to the fact 
that a larger amount of pulp may be used for each individual assay, 


thus increasing the accuracy of the results. The copper is eliminated — 4 


as it is in the scorification assay by the solution of its oxide in the basic 
lead oxide slag. The assay thus combines the advantages of the scori- 
fication with those of the crucible assay. 
Perkins ' has made a careful study of this process, and calls atten- 
tion to the fact that the litharge used must be in proportion to the _ 
1 The Litharge method of Assaying Copper Bearing Ores and Products, and the a 
Method of Calculating Charges. W. G. Perkins, T. A. I. M. E., 31 p. 913. 





is 3 


amount of copper and other impurities in the ore. The amounts he 
uses are very large (from 137 to 300 parts PbO to | part Cu), and make 
the method an expensive one. Others have reduced this amount 
considerably, and still manage to get buttons which will cupel. 

The Slag. The slag should be decidedly basic, for if we combine 
the litharge with large amounts of silica-and borax, it will no longer 
retain its power of holding the copper in solution. A small amount 
of silica is necessary to prevent to some extent the action of the 
litharge upon the crucible. One part of silica to from 15 to 20 parts 
of litharge is generally allowed in the charge. Borax should be en- 
tirely omitted as it acts to decrease the copper holding capacity of 
the slag, and also causes boiling of the charge. Perkins states that 
the best results are obtained with a slag which exhibits when cooled 
and broken a somewhat glassy exterior gradually passing to litharge 
like crystals towards the center. The amount of crystallization which 
takes place is, of course, a function of the rate of cooling and will 
depend among other things upon the size of the charge, the tempera- 
ture of the charge when poured, and of the mold, so that too much 
weight should not be given to the above. ‘The slag should however 
be crystalline resembling litharge, and if dull or glassy throughout, 
indicates the presence of too much acid for a good elimination of 


— copper. 





Conduct of the Assay. On account of the very corrosive action of 
the litharge slag it is especially necessary that the fusion be made 
rapidly. The muffle should be hot to start (1000° to 1100° C.), 
the hotter the better, and the fusion should be finished in from 20 
to 30 minutes. This not only preserves the crucibles, but also as a 
necessary sequel prevents the slag from becoming charged with silica 
and thus forcing the copper into the button. The slag melts at a low 
- temperature and a very high finishing temperature is not necessary. 
With a quick fusion there is less chance for oxidation of lead with the 
consequent reduction of too small a lead button. 

For the best work the hole in the back of the muffle should be 

stopped up, and a reducing atmosphere maintained in the muffle. 

This may be accomplished by filling the mouth of the muffle with 
charcoal or coke, or by distributing a few crucibles part full of soft 
coal throughout the charge and using a tight-fitting door. If this 
precaution is not observed part of the silver will be oxidized and lost 
in the slag. 

The following charges. kindly furnished by the Boston and Mon- 
tana Reduction Department of the Anaconda Copper Mining Com- 
pany, Great Falls, Montana are recommiended for these ores. 


116 



































TABLE XXIII. CHARGES FOR COPPER BEARING MATERIAL. 
Approximate Charge for Silver Charge for Gold 
Material Analysis (In 20 gram crucible) | (In 30 gram crucible) 
Cu 9%-15% Sample 1 A.T. | Sample Lage 
SiO 15% 23% Soda 20 grams | Soda 30 grams 
FeO 33 %—-40% Litharge 100 “ Litharge 150 ‘“ 
Concts. S 33 %—40% Silica yee Silica olen 
Ag 3 02.—5 OZ. NitsE) Lo 20s Niter 40-60 “ 
Au 0.0150z.-—0.0250z.| Cover mixture Cover mixture 
Cu 30%—-45% Sample SAE: Sample 2 A. T. 
Fe 40 %-30% Soda 18 grams | Soda 25 grams 
Matte S 30%-27% Litharge 100 “ Litharge 200 “ 
Ag 10 0z.-18 oz. Silica Te Silica IDivke 
Au 0.07 oz.-0.11 oz. | Niter ae Niter 1S 
Cover mixture Cover mixture 
Cu 45%-60% Sample 1A. T. Sample A 
Fe 30Z-15% Soda 18 grams -| Soda 25 grams 
Matte 8 27%-24% Litharge 125 “ Litharge 240 “ 
Ag 15 0z.-25 oz. Silica (jee Silica 1a 
Au 0.10 oz.—0.14 oz. | Niter 7 hee Niter 14505 
Cover mixture Cover mixture 








— 











Assay of Antimonial Gold Ores. The niter method-is universally 
recognized as being the best method for the sulphide ores of antimony. - 
Considerable litharge is necessary to keep the antimony out of the 
lead button. The following charge is recommended by two English 


authorities: — 
Ore 5A. T. Litharge 100-120 grams 
Na,CO; 10-20 grams Niter LO ee 
Borax-glass 5-10 “ Silica 10 Say 


A preliminary assay to determine the reducing power is of course 
necessary. The above charge will be found to correspond almost 
exactly with our standard for sulphide ores, with litharge according 
to Lodge’s rule. 

George T. Holloway in discussing this method recommended using 
a much larger proportion of soda in the charge, 1.e., three times as 
much as stibnite, in order to aid in the retention of the antimony in 
the slag as a sodium antimoniate. 


Assay of Auriferous Tinstone. C. O. Bannister’ finds a crucible 
assay with the following charge to be the most satisfactory 
method :— 


' William Kitto, Tr. Inst. of Min. & Met., 16 p. 89. 
' William Smith, Tr. Inst. of Min. & Met., 9 p. 332. 
2 Trans. Inst. of Min. and Met. (London) 15 p. 513. 








Shoe ee ee tess 25 grams 
Memmmacarpondte..;...2..2i7..2 > 40° “ 
eh). ho Sok ee, LOR? 
eM are So 4 ober ee 
mmr at Gg a eS Lo yee 


In this method the tin is converted into a fusible sodium stannate. 
The author found no tin reduced during the fusion as shown by the 
button cupelling without difficulty. In all ores carrying over 1 oz. 
of gold per ton, the slags were cleaned by a second fusion with 10 
grams of soda, 30 grams of red lead and 1.5 grams of charcoal. 

Various other methods of assay were tested but none were as satis- 
factory as this. 

Corrected Assays. In the assay of high-grade ores and bullion 
it is often desirable to make a correction for the inevitable slag and 
cupel losses. This is done in one of two ways: either by the use of 
a “check” or synthetic assay or by assaying the slags and cupels re- 
sulting from the original or commercial assays. 

In correcting by a ‘“‘check’”’ assay a preliminary assay is first. made 
and then an amount of proof silver or gold, or both, approximately 
equivalent to the amount present in the sample, is weighed out and 
made up to approximately the composition of the sample by the ad- 
dition of base metal, etc. The check thus made is assayed in the 
same furnace parallel with the real assay. Whatever loss the known 
amounts of precious metal in the check sustain is added to the weight 
of metal obtained from the sample as a correction, the sum being sup- 
posed to represent the actual metal present in the sample. This 
method of correction is always used in the assay of gold and other 
precious metal bullions, and is sometimes used in the assay of high- 
grade ores. A more detailed description of the method will be found 
in the chapter on the assay of bullion. This method when properly 
applied is the better and gives a very close approximation to the actual 
precious metal contents of a sample. 

In the case of rich ores and furnace products other than bullion, 
a correction is usually made by assaying the slags and cupels result- 
ing from the original assay. The metals thus recovered are added 
as a correction to the weight first obtained. This method, while 
approximating the actual contents of an ore, may occasionally give 
results a little too high, for although gold and silver lost by volatiliza- 
tion is not recovered and the corrections themselves must invariably | 
suffer a second slag and cupel loss, yet on the other hand, the cupelled 
metal from both the first and second operations is not pure and may 
retain enough lead and occasionally other impurities from the ore 


118 


to more than offset the above small losses. The results of assays cor- 
rected by this method are evidently somewhat uncertain, but are 
nevertheless much nearer to the real silver content than are the re- 
sults of the uncorrected or ordinary commercial assay. 

Smelter contracts are almost invariably still written on the basis 
of the ordinary or uncorrected assay and when the corrected assay 
is made the basis of settlement, a deduction is made amounting to 
the average correction. This amounted to 1.1 per cent in the case 
of certain Cobalt ores. 

When a corrected assay is to be made it is well to use a Portland 
cement or magnesia cupel for the first cupellation as these materials 
-are easier to flux than bone-ash. 

To assay a Portland cement cupel the following charge, a sesqui- 
silicate, is found to give satisfactory results :— 

Cupel (45 grams of cement). 

Na2zCO;—45 grams, 

Borax-glass—21 grams. 

Litharge—Dependent on size of original button, 
to make a total of 75 grams. 

Argols (R. P. 10) 3.2 grams. 

Silica—32 grams. 

To assay a magnesia cupel, good results are obtained by adding 
soda equal to the original weight of the cupel and litharge to make a 
total of 30 grams more than the original weight of the cupel (allow- 
ing for litharge in the cupel). Compute the silica necessary to make 
a sesqui-silicate with magnesia, soda and active litharge, and add 
two-thirds of this weight of silica and substitute for the other third 
twice its weight of borax-glass. ) 

To assay a bone-ash cupel, first remove and reject the unsaturated 
part of the cupel in order to have as little of this refractory material 
as possible to deal with. The saturated part will be about 50 per 
cent bone-ash and 50 per cent litharge. Grind to 80 mesh and clean — 
the bucking board or machine by grinding 10 or 15 grams of 10 mesh q 
silica. This should be reserved and added to the charge. To assay, 
add a weight of soda equal to the weight of saturated cupel material, — 
two-thirds as much borax-glass, 25 grams of litharge plus enough © 
more to make a total equal to the weight of saturated cupel material, — 
silica one-third as much cupei material, reduane agent for a es gram 
lead button. For example:— : 


Cupel material}. : 45 grams Litharge 47% grams 
od askOave tok aie ee Argols (R. P.10) 2.2 “ 
. Borax-glass . : pOaia. Silica (from clean- 


ing board) tbo 








CHAPTER X. 
THE ASSAY OF BULLION. 


Bullion from an assayer’s point of view is an alloy containing enough 
of the precious metals to pay for parting. 

The different bullions are usually named to correspond with their 
major components, for instance, copper bullion an alloy of copper 
with small amounts of other impurities, as well as some gold and silver. 
In the same way we have lead, silver and gold bullions. Doré bullion 
is silver bullion containing gold. The term base bullion is used in 
two different senses. According to the lead smelters definition base 
bullion is argentiferous lead, usually the product of the lead blast 
furnace; according to the mints and refiners definition it is bullion 
containing from 10 to 60 per cent of silver, usually some gold, and a 
large percentage of base metals particularly copper, lead, zine and 
antimony. Fine gold bars are those which are free from silver and 
sufficiently free from other impurities to make them fit for coinage 
and use in the arts usually 990 to 999 fine. 

The results of lead and copper bullion assays are reported | In ounces 
per ton as in the case of ore assays, but in the assay of silver, gold and 
doré bullions the results are reported in “‘fineness,” i.e., so many 
parts of silver or gold in one thousand parts of bullion. Thus sterling 
silver is 925 parts fine, that is to say, it is 92.5 per cent silver, 


Weights. In assaying gold, silver and dore bullion, a special 
set of weights called gold assay weights are used. This is termed 
the ‘“‘millime”’ system, and the unit one millime weighs 0.5 milligram, 
and therefore the 1000 millime weight equals 0.5 grams. Ordinary 
weights in the gram system may be used but as 0.5 gram is the quan- 
tity of bullion commonly taken for assay the use of the millime system 
saves computation in obtaining the fineness. 


Sampling Bullion. 


Bullion may be sampled either in the molten or in the solid condi- 
tion. When it may be melted and kept free from dross the dip or 
ladle sample is usually the more accurate method. As the weight, 
as well as the assay of the bullion must be known in order to value 
it, the sampling of large lots of bullion by the dip sample method 


120 


often presents difficulties owing to changes in weight or purity in 
the considerable length of time necessary for pouring. Again it is 
not always convenient to melt a lot of bullion to obtain a sample, 
and other means must be found. Sampling solid bullion by punch- 
ing, boring, sawing or chipping, under certain conditions, may be 
made to yield good results. Lead bullion is usually sampled by punch- 
ing one or more holes in each bar, and combining and melting the 
punchings. Copper bullion is now generally cast in the form of slabs 
or anodes, and these are drilled. 

Sampling Molten Bullion. The most satisfactory method of 
sampling bullion is to melt the whole in a suitable vessel, stir 
thoroughly with a graphite rod or iron bar to mix and then immediately 
before pouring, ladle out a small amount and granulate it by pouring 
into a pailof water. If these operations are correctly performed there is 
no chance for segregation, and each particle of the granulated metal 
should be a true representative of the whole. If a granulated sample 
is not desired, a ladleful of the mixed molten metal may be poured 
into a thick-walled flat mold so that it chills almost instantly, and a 
drill or saw sample may be taken from this. When a ladle sample is 
taken, the ladle must be so hot as not to allow the forming of any 
solidified metal or “‘sculls’’ as this would interfere with the homogeneity 
of the sample. This method of sampling is most satisfactory on 
bullons which do not oxidize or form dross on melting, as this of 
course, adds a complication hard to allow for. 

Sampling Solid Bullion. The principal difficulty in the sampling 
of bullion in the form of bars or ingots is caused by the segregation 
of the various metals in cooling. If it were, possible to cool a bar 
instantly, segregation would be prevented, and a chip or boring taken 
from any part would be representative. As instant cooling is im- 
possible, the sampling of bars of the ordinary dimensions becomes a 
difficult problem. As the result of a careful study of this problem 
Keller' has concluded that it is almost impossible to obtain samples 
of satisfactory accuracy from bars or pigs of the usual dimensions. 
To eliminate the difficulties of sampling from a bar he recommends 
casting the metal in the form of a thin plate. Of course some con- 
centration would take place here also, but as the plate would solidify 
so much faster than the same metal cast in a bar or ingot this factor 
would have less weight. Owing to the fact that concentration takes 
place from or toward every surface, we will have all around the plate 
a zone not wider than the thickness of the plate where concentration 
has taken place both horizontally and vertically, but which should 

1 'T. A. I. M. E,, 27, p. 106. 








121 


of itself be a sample of the whole. In the part of the plate enclosed 
by this zone we have concentration in the vertical direction only. 
If we drill or punch through this part of the plate we should obtain 
a correct sample of the whole. Keller cites experiments to prove the 
above theory. 


Some typical methods of sampling lead and copper bullion follow. 


Sampling Lead Bullion. Lead bullion is sampled both in 
the liquid and in the solid state. In either case it is now customary 
to transfer the lead from the blast furnace either into a reverberatory 
furnace or into large kettles holding 20 to 30 tons. Here it is purified 
either by liquation, or by cooling to a little above the melting point of 
pure lead. By doing this, a large part of the impurities which are 
held in solution by the superheated lead are separated out as a dross 
which is carefully removed by skimming. The remaining lead, which 
is now in a better condition to sample, is drawn off by means of a 
syphon and cast into bars of about 100 pounds. 

In taking a dip sample a small ladleful is taken at regular intervals 
from the stream coming from the syphon. These individual samples 
are carefully remelted at a dark red heat in a graphite crucible, the 
melt is well stirred and cast in a heavy-walled shallow mold, making 
a cake about 10’’ long, 5’’ wide and 4’’ thick. This cools so quickly 
that there is little or no chance for segregation. The final assay 
samples are taken from this cake either by sawing and taking the 
sawdust, or by boring entirely through the slab in a number of places, 
and taking the borings, or by cutting out four or more 3 A. T. pieces 
from different parts of the bar and using these directly. 

In sampling solid lead bullion the bars are sampled by means of a 
heavy punch which takes a cylindrical sample about 2’’ long and 
1/8’’ in diameter. There are naturally a number of different systems 
but the most common method is to place five bars side by side and 
face up, and punch a hole in each extending half-way through. Each 
bar is punched in a different place and in such a way that the holes 
make a diagonal across the five bars. The bars are then turned over 
and another sample is taken from each along the opposite diagonal. 
Usually one carload of about 20 or 30 tons is sampled as one lot. The 


-punchings from such a lot, weighing from 8 to 15 pounds are melted 


in a graphite crucible and cast into a flat bar, from which the final 
assay samples are taken by sawing, drilling or cutting. 


Sampling Copper Bullion. The sampling of copper bullion may 
be classified into smelter methods, and refinery methods. The bullion 
is quite universally cast in the form of anodes at the smelter, and 


] 


122 


shipped to the refinery in this form. This renders remelting at the 
refinery unnecessary, and the result is that the refiners sample the 
solid bullion by drilling. The smelters, having the bullion in the 
molten condition, generally sample it in this condition on account of 
the greater ease and less expense. 

Probably the most satisfactory smelter method of sampling is the 
“splash shot method”’, which consists in shotting into water a small 
portion of the molten stream of copper as it flows from the refining 
furnace by “batting” the stream with a wet stick. This operation 
is repeated at uniform intervals during the pouring, the amount taken 
each time being kept about the same. The samples are dried and 
dirt and pieces of burned wood are removed. All material over four 
mesh and under 10 mesh is rejected, and the remainder taken as the © 
sample. This method when properly carried out gives results which 
check within practical limits with the drill sample of the anodes taken 
at the refinery. 

Another method which is used to some extent for sampling molten 
copper bullion is known as the ‘“‘ladle-shot method.” This consists 
in taking a ladleful from the furnace or from the stream of the casting 
machine and shotting it by pouring over a wooden paddle into water. 
In this method at least three ladlefuls are taken, one near the begin- 
ning, one at the middle, and one near the end of the pour. The shots 
are treated in the same manner as before. This method is not thought 
so well of as the previous one on account of segregation toward or 
from the “ sculls ”’ which are left in the ladles. 

Instead of shotting and taking the shot for the final sample, W. H. 
Howard of Garfield, Utah, recommends ladling into a flat dise. 
This ‘‘pie sample” is sawed radially a number of times, and the saw- 
dust used for the final sample. 

The following description of the method of sampling anodes at 
Perth Amboy, N.J. is typical of refinery methods of sampling and is 
the method developed by Dr. Edward Keller. The copper is re- 
ceived in the form of anodes 36”’ long, 28’’ wide and 2”’ thick. These 
are carefully swept to remove foreign matter, and then drilled with 
a 0.5’’ drill completely through the anode, all of the drillings being 
carefully saved. A 99-hole template is used to locate the holes which 
are spaced 3 1/16’’ center to center, and the outside row is approxi- 
mately 23’’ from the edge of the anode. The holes of the template 
are used in continuous order, one hole to the anode. 

For very rich anodes some refiners use a template having as many | 
as 240 holes, but it seems doubtful if this arrangement of spacing a 
single hole in each anode will yield any better sample. 





——— tan 





123 


With low-grade, uniform bullion every fourth anode only is drilled. 
A 30 ton lot of anodes in which each one is drilled will yield 6 or 8 
pounds of drillings, which are ground in a drug-mill fitted with man- 
ganese steel plates and reduced by quartering to about 2 pounds. 
This sample is reground until it will all pass a 16-mesh screen and is 
then divided into the sample packages. 


The Assay of Lead Bullion. 


A description of the cupellation assay of lead bullion has already 
been given in the chapter on cupellation. In smelter control work 
the assay is usually made in quadruplicate. If the bullion contains 
sufficient copper, arsenic, antimony, tin or other base metals to 
influence the results of the cupellation assay, three or four portions 
of 0.5 or 1.0 A. T. are scorified with the addition of lead until the 
impurities are eliminated, when the resultant buttons are cupelled. 

Correction for Cupel Loss. In some instances the slags and cupels 
are re-assayed and the weight of the gold and silver found is added 
to that obtained from the first cupellation. There is no fixed custom 
as yet regarding the use of corrected assays. In most of the custom 
smelters, the uncorrected assay is used as the basic of settlement; 
but some of the large concerns who have their own refineries are using 
the corrected assay in their inter-plant business. 


The Assay of Copper Bullion. 


Copper bullion may be assayed by the scorification, crucible or 
by a combination of wet and fire methods. In the combination method 
the bullion is treated with sulphuric or nitric acid which dissolved 
the copper and the silver and leaves the gold. The silver is pre- 
_cipitated by suitable reagents and filtered off together with the gold. 
The filter paper and contents are put into a scorifier or crucible with 
reagents and the assay finished by fire methods. 

The scorification method is generally accepted as standard for gold, 
and many smelter contracts state that ‘gold shall be determined by 
the all fire method or its equivalent.’’ The mercury-sulphuric acid 
combination method on many bullions gives gold results equal to 
the scorification. The silver results obtained by the scorification 
method are open to suspicion owing to the considerable slag and 
~ cupellation losses, and the doubt concerning the purity of the buttons, 
which often contain noticeable amounts of lead and copper. 
~The crucible method has not as yet come into common use for the 
determination of gold and silver in copper bullion, but according to 


124 


Perkins! it gives gold results equal to the “all-scorification‘’ method. 
In smelter practice, silver in copper bullion is determined usually 


by the nitric acid combination method, sometimes by the mercury- 


sulphuric acid combination method. This later method tends to 
give high silver results, owing to the incomplete solution of the copper 
in the acid, and the possibility of some copper being retained in the 
silver bead. 

The nitric acid combination method is recognized as giving low 
results in gold. Van Liew’ attributes this to the solution of the gold 
in the mixture of nitrous and nitric acids present. He found a decided 
loss (33.7%) of gold on treating gold leaf with a mixture of nitrous 
and nitric acids for two and a half hours. He gives a method of slow 
solution in cold dilute acid which reduces this loss to a minimum. 


The Scorification Method. The following method commonly 
referred to as the “‘all fire’’ method is a modification kindly supplied 
by Mr. H. D. Greenwood, Chief Chemist for the United States Metals 
Refining Co., Chrome, N. J. 

Sample down the finely ground bullion on a split sampler in such a 
way as to obtain a sample of about 1 A. T., which will include the 
proper proportion of finer and coarser parts of the borings. This 
sampling must be conducted carefully, as the precious metal contents 
of the finer parts differs somewhat from that of the coarser portion 
of the sample. Portions “dipped” from the sample bottle or from the 
sample spread out on paper are likely to contain undue amounts of 
coarse or fine. 

Weigh out 4 portions of copper borings of } assay ton each, mix with 
50 grams test lead, put in 38-inch Bartlett scorifiers, cover with 40 
grams test lead and add about 1 gram SiOs. Scorify hot, heating 
at finish so as to pour properly. Add test lead to make weight of 
button plus test lead equal to 70 grams, add 1 gram SiO, and scorify 
rather cool. Pour, make up to 60 grams with test lead, adding 1 
gram SiO: and scorify. 

Combine the buttons two and two, and make up each lot to 85 
grams with test lead, adding | gram SiO» and scorify very cool. Make 
up button to 70 grams by adding test lead, add 1 gram SiO. and 
scorify for the fifth time. The buttons should be free from slag and 
weigh 14 grams. 

Cupel at a temperature to feather nicely, and raise heat at finish. 
Cupels to be made of 60 mesh bone-ash, and to be of medium hard- 
ness. 


PPA. LANA Boo ppuonl: 
2E. & M. J. 69, pp. 496, et seq. 





a my cafe eee 


Pee | wee oe 





125 


Weigh the bead and part as usual. Dry, anneal and weigh. The 
two results should check within .02 oz. per ton, and the average 
figure is to be reported. If the silver contents of the bullion is low, 
add enough fine silver to the copper borings before the first scorifica- 
tion to make the total silver in the mixture equal to about 8 times the 
amount of gold. 


The Crucible Method. The crucible method for gold and silver 
in copper bull.on was first described by Perkins! and as described by 
him showed no great advantage over the scorification method as to 
saving in time or cost of materials, or increased furnace capacity. 
The following modified procedure requires about one-third of the 
materials, time and furnace capacity as that described by Perkins, 
and yet gives buttons sufficiently free from copper so that they may 
be cupelled direct. 

Sample down the finely ground bullion to about + A. T. and adjust 
the weight of the sampled portion to exactly + A.T. Place in a 
20 gram crucible and mix with it 1.2 grams of powdered sulphur. 
Cover this with a mixture of 15 grams of sodium carbonate, 240 
grams of litharge, and 8 grams of silica; but do not mix with the sul- 
phur and copper which should be allowed to remain in the bottom of 
the crucible. Cover with salt or flux mixture and place in a hut muflie 
so that the charge will begin to melt in 6 or 8 minutes. ‘lhe fusions 
should be quiet and ready to pour in 25 or 30 minutes. 

If a salt cover is used the lead buttons should weigh abeut 32 grams, 
if a flux cover is used they may be somewhat smaller. With a properly 
conducted assay the buttons are soft enough for direct cupellation; 
but the cupels are quite green. If the assayer prefers, the buttons 
may be made up to 50 or 60 grams with test lead and scorified in a 
3 inch scorifier to further eliminate the copper. After cupellation 
_ the buttons are weighed and parted as usual. It is well to do four 
fusions, and to combine the buttons two and two for parting. 


Remarks. As soon as the sulphur melts (115° C.) it combines with 
the copper to form a matte which is decomposed and most of its copper 
oxidized and slagged by the litharge of the charge. The fusions melt 
down very quietly almost without boiling, and with a short period 
of fusion the crucibles are not badly attacked. The final temperature 
need not be higher than a good bright red or full yellow. The slag 
is heavy but very fluid and should not contain any lead shots. 

The method gives results in gold equal to the scorification method; 


1 An “All-Fire’’ Method for the Assay of Gold and Silver in Blister Copper. 
W. G. Perkins, T. A. I. M. E., 33 pp. 670. 


126 


but like any method using high litharge, the silver is apue to be some- 
what low. : 


Mercury—Sulphuric Acid Method. Sample down the’ finely 
ground bullion on a split sampler in such a way as to obtain a sample 
of about 1 A. T., which will include a proper proportion of the finer 
and coarser parts of the borings. This sampling must be conducted 
carefully as the precious metal contents of the finer parts differs some- 
what from that of the coarser portion of the sample. Portions 
“dipped”’ from the sample bottle or from the sample spread out on 
paper are likely to contain undue amounts of the coarse or fine. 

Adjust the weight of the sampled portion to exactly 1 A. T. and 
transfer it to a No. 5 beaker (capacity about 750 ¢c.c.). The beaker 
should have a watch glass cover. 

Treat the sample with 10 c.c. of mercury solution and shake the 
beaker until the copper is thoroughly amalgamated; then add 80 
c.c. of strong sulphuric acid, place the beaker on a hot plate and 
boil until the copper is dissolved. This requires about twenty 
- minutes. 

Remove the beaker and allow it to cool. The contents will be a 
semi-liquid sludge. When cool, add about 100 ¢.c. of cold water and 
mix; then add about 450 c.c. of boiling water and stir until the copper 
Bilphite dissolves. 

Bring to boiling and add 4 to 6 c.c. of salt soins 1 c.c. equivalent 
to 50 mg. Ag. Remove from the hot plate and add 10 c.c. of a 10 
per cent solution of lead acetate. Stir well, allow to settle and filter 
at once through double filter papers (124 or 15 c.m.) washing the 
beaker with hot water. Finally wipe the inside with filter paper and 
add it to the filter. Thorough washing of the filter is not necessary. 

Transfer the wet filter and its contents to a 24’’ scorifier which 
has been glazed on the inside by melting litharge in it and pouring 
away the excess. 

Burn off the filter paper at a low temperature—best in a closed oven 
which may be heated to, say 175°C. This chars the paper slowly 
without danger of loss of silver. 

When the paper is consumed, add 30 grams of test lead and scorify; 
pour so as to obtain a 12 gram button, cupel as usual to produce feather 
litharge, weigh the gold and silver bead and part with dilute nitric 
acid. 

The mercury solution mentioned above is made by dissolving 100 
grams of pure mercury in nitric acid, and diluting to 500 c.c. From 
this stock solution take 50 c.c. and dilute it to 1 liter, for the working 





| 
: 
: 
% 
: 
: 
: 





127 


solution—10 c.c. of the latter will serve for each assay-ton of copper 
bullion. 

The object of adding mercury is to secure an easy solution of the 
copper in sulphuric acid. If the copper is treated directly without 
previous amalgamation, it is very difficult to dissolve it in sulphuric 
acid. In fact a considerable portion of it will remain insoluble, 
partly in the form of sulphide of copper. If the copper be amal- 
gamated on the other hand, solution proceeds smoothly until prac- 
tically all of the copper is dissolved. When the bullion is low in 
precious metals, say less than 50 ounces per ton, no silver dissolves 
in the sulphuric acid. No gold dissolves whatever the grade. If the 
bullion is very rich in silver a little of it may dissolve in the acid. 

The assays should be made in duplicate or triplicate, and the aver- 
age results reported. Differences on silver seldom exceed 0.2’ ozs.; 
on gold the results are usually exactly the same. The sulphuric 
acid used should be chemically pure and full strength (1.84 sp. gr.). 


Nitric Acid Combination Method.’ Sample down the finely 
ground bullion on a split sampler in such a way as to obtain a sample 
about 1 A. T., which will include the proper proportion of the finer 
and coarser parts of the borings. This sampling must be conducted 
carefully as the precious metal contents of the finer parts differ 
somewhat from that of the coarser portion of the sample. Portions 
“dipped” from the sample bottle or from the sample spread out on 
paper are likely to contain undue amounts of coarse or fine. 

Weigh out two portions of copper borings of 1 A. T. each, and carry 
the assay through on each portion as follows:— _ 

Place in a No. 5 beaker, add 100 c.c. of distilled water and 90 c.c. 
HNOs, sp. gr. 1.42, the latter being added in portions of 30 c.c. each 
at intervals of about 1 hour. When all is in solution precipitate a 
small amount of silver chloride with salt solution in order to collect 
the gold, filter through double filter papers and wash the filter papers 
free from copper. To the filtrate add the calculated amount of salt 
solution to precipitate all the silver and a slight excess, measuring the 
solution with a burette and varying the amount added with the rich- 
ness of the bullion. Allow to stand over night after stirring well. 
Filter the silver chloride through double papers, wash papers free 
from copper, then sprinkle 5 grams of test lead in the filter paper and 
fold into a 234 inch Bartlett shape scorifier, the bottom of which is 
lined with sheet lead. To this add also the filter papers containing, 


“1 Procedure kindly supplied by Mr. D. H. Greenwood, Chief Chemist, for the 
United States Metals Refining Company, Chrome, N. J. 


128 


the gold. Dry and ignite the filter papers carefully, cover with 
35 grams of test lead, a little borax-glass, and scorify at a low heat 
so that the resultant button will weigh about 12 grams. Cupels — 
should be feathered nicely. Cupel to be made of 60 mesh bone-ash 
and to be of medium hardness. Weigh the bead and part. Anneal 
and weigh the gold. The two results on gold should check within 
0.02 oz. per ton, and the silver within 1 per cent. 


The Assay of Dore Bullion. 
This method is the one generally adopted by assayers in this coun- 


try, and may also be used for the assay of silver bullion. A better 
method for the accurate determination of silver in doré or silver bullion 


is probably the Gay-Lussac or salt titration, also known as the mint — 


method. This later method requires considerable equipment and 
preparation and for this reason the occasional assay is more easily 
done by fire methods. 

The Check. In order to correct for the inevitable losses in cupelling 
as well as for any other errors in the assay, silver, doré, and gold 
bullions are always run with a check. This check or “proof center” 
is a synthetic sample made up of known weights of pure silver, gold 
and copper to approximate as closely as possible the composition of 
the bullion to be assayed. It is cupelled at the same time and under 
the same conditions as the regular assays, and whatever gain or loss 
it suffers is added as a correction to the regular assay. To obtain 
data to make up the check a preliminary assay is made to determine — 
the approximate composition of the bullion. 

Preliminary Assay. A sample of 500 mgs. of bullion or as near this 
amount as possible is weighed out on the assay balance, and the exact 
weight recorded. This is compactly wrapped in 6 or 8 grams of lead 
foil and cupelled in a small cupel with feather crystals of litharge. 
The cupel should be pushed back in the muffle for the last two or 
three minutes to insure the removal of the last of the lead. After 
the play of colors has ceased it should be drawn toward the front of 
the muffle and then covered with a very hot cupel to prevent sprout- 
ing. It is then removed gradually from the muffle and when cool 
is cleaned, weighed and parted in the ordinary manner. The gold 
will require more than the ordinary amount of washing on account 
of the large quantity of silver present. 


If the cupelling has been properly conducted it will be fair to assume 


a loss of one per cent of silver in determining the approximate silver. 
The weight of gold may be taken as approximately correct. The 





ORO OT ee a | a ee ae ee oe 





129 


~ sum of the weights of approximate gold and silver is subtracted 


from the weight of bullion taken to obtain the amount of base metal. 
This will usually be copper, but whatever it is the assayer should be 
able to determine by the appearance of the bullion and the cupel. 


Final Assay. Two portions of approximately 500 mgs. are weighed 
accurately and wrapped in the proper amount of lead foil as shown 
by the following table which assumes the impurity to be copper. 


TABLE XXIV. LEAD RATIO IN CUPELLATION. 




















Fineness of Au. + Ag. | Wt. of Lead Fineness Au. + Ag. Wt. of Lead 
950 5 gms. 750 llgms 
900 7 Os ' 700 12 66 
850 8 650 ist: 
800 PO eae s 600 15a 








A check is made up with C. P. silver and proof gold equal to the 
approximate silver and gold found by the preliminary assay and the 
necessary amount of copper or other base metal. These are wrapped 


’ up in the same amount of sheet lead as was used for the bullion. The 


lead for these assays is best cut into equal sized rectangles with pro- 
portions approximately 13’ X 23’’, and twisted intu (the whape of 
little cornucopias with the bottoms folded up. The bullion aid metals 
going to make up the check are transferred to these directly from the 
seale pans after which they are folded over and made into compact 
bundles. 

The cupels are placed in a row across the muffle and when they are 
hot the buttons are dropped quickly into them with the check in the 


middle. They should be cupelled at a low temperature so that plenti- 


ful crystals of litharge are obtained all around the buttons, but toward 
the end the temperature should be increased to insure driving off. 
the last of the lead. 

The buttons are cleaned, weighed, parted and the gold weighed. 
The per cent loss of gold and silver is determined and a corresponding 
correction made to the weights of gold and silver found. From these 
figures the fineness in both gold and silver is determined. The gold 
should check within 0.1 part and the silver within 0.5 parts. 

Notes. 1. When the doré contains antimony weigh the samples into 2.5’ 
scorifiers with 30 grams of test lead. The proofs are made up according to the 
preliminary assay. All are scorified in the same muffle at the same time. Pour 
and hammer the lead buttons into a cube. Should the weight of these lead buttons 


vary over a gram, make up to the same weight with sheet-test-lead, cupel and 


part as usual. / ; | 
2. When the doré contains bismuth, selenium or tellurium, three 3 gram 


130 


portions are weighed out into 24’’ scorifiers with forty grams of test lead, scorified 
and the lead buttons flattened out into a sheet about 3 inches square. This sheet 
of lead is dissolved in about 200 c.c., of dil. HNOs (1-3) and boiled to expel all red 
fumes. Dilute to 400 c.c., filter through triple folded 15 em. filter, washing the pre-_ 
cipitate only once. To the filtrate is added sufficient NaCl solution to precipitate 
all the silver. Heat to boiling and allow to stand over night. Filter through 15 em. 
filter washing the precipitate only once. Place the two filter papers in a 23’" lead 
lined scorifier, dry in an oven, burn, then cover with 30 grams of test lead and 
scorify. Open the scorification at a rather high heat, continuing with a gradually 
falling temperature. When the scorifiers have entirely closed over, close the muffle 
door, raise the heat and pour; then treat exactly as in No. 1. 

If the silver fineness of the doré is not three or more times greater than the pole 
fineness, another set of assays must be run with the addition of proof silver at the 
weighing out of the doré. ; 


U. S. Mint Assay of Gold Bullion. 


Melting. Every lot of bullion or dust received at any U. S. Assay 
Office or Mint is immediately weighed and given a number. It is 
then sent to the melting room. Here it is melted in a graphite crucible 
with borax and soda and cast into a bar. Usually no attempt is 
made to refine it unless it is very impure. Occasionally, in the case 
of very impure bullions, a small dip sample is taken and granulated, 
but in general the whole melt is cast and sampled as noted below. 
The slag is poured with the bar and when solid is ground, panned and - 
the recovered prills are dried, weighed and allowed for in computing 
the value of the bar. | 

Sampling. After the bar is cleaned of slag it is dried, weighed and 
numbered and from diagonally opposite corners two samples of 3 or 
4 grams each are chipped. These are flattened with a heavy hammer, 
annealed and rolled into sheets thin enough to be easily cut with 
shears. The use of the shears can only be learned by practice, but 
assayers become very skillful after a time so that it is no unusual thing 
to see a bullion assayer weigh out five samples in almost as many 
minutes. 

Preliminary Assay. Assay for Bases. To determine the ap- 
proximate composition of the bullion a preliminary assay is made. 
A sample of 1000 gold weights (500 mgs.) is weighed out, wrapped in 
five grams of lead foil and cupelled. The weight of the bullion taken, 
less the weight of the button obtained gives the base metals. 

The button now consists of gold and silver, the approximate relative 
proportions of which must be determined. This may be done by 
adding silver, cupelling and parting or by touchstone. This later 
method is used at the Government Assay Offices and Mints. The — 
touchstone consists of a piece of black jasper on which the sample is 
rubbed and the mark compared with marks made with alloy slips 
(needles) of known compositoin. The needles range from 500. to 








131 


1000 fine and are 20 points apart. This gives the fineness within 
2 per cent which is close enough to show how much silver to add to 
inquart the main assay and to make up the check or proof center. 

Final Assay. The final assay is usually made by two assayers each 
working on one of the chipped samples. In the case of a small bar 
each makes one assay, while in the case of a large bar each assayer 
makes two or more assays. The balances used for the assay are 
usually adjusted so that a deviation of the needle of one division on 
the ivory scale amounts to some simple fraction of the weights used. 
Thus at one assay office a deviation of the swing of one division on 
the ivory scale amounts to 0.1 mg. = 0.2 gold weights. With this 
adjustment it is not necessary to make so many trials with the rider 
to get the final weight, nor is it necessary to weigh out exactly an even 
half gram of bullion for the assay. Instead we weigh out 1000 = 3 
divisions on the ivory scale, record the difference, and make a cor- 
responding correction when the gold cornet is weighed. 

As stated above the weight of bullion taken for each assay is 
1000 (500 mg.). To this is added sufficient silver to make the ratio 
of silver to gold 2 to 1, and the whole is wrapped up in 5 or 6 grams 
of lead foil. The lead foil pieces are all cut to exact size, about 14’’ 
23’’, and rolled up into cornucopia shape with the bottom pinched in. 
The bullion is poured directly into these from the scale pan. ‘I'he 
silver is added in the form of discs made for convenience into 4 or 5 
different sizes. These discs are punched out of sheets carefully rolled 
to gauge so that the punchings will weigh exactly even tens and hun- 
dreds in the gold weight system. If the bullion contains no copper 
it is advisable to add about 30 parts gold weight (15 mg.). This cop- 
per may be alloyed with the silver used for parting. 

One or more proofs of pure gold weighing usually 900 (0.450 grams) 
are also weighed and made up to the 2 to | ratio and copper added to 
approximate that in the bullion. These are wrapped in the same 
quantity of lead foil as the bullion, and one or more are run in each row 
of cupels in the muffle. The lead packets are pressed into spherical 
shape by pliers specially designed for the purpose. 

The lead packets are put in order as prepared in the numbered 
compartments of a wooden tray and taken to the furnace room where 
they are cupelled in a rather hot muffle. The cupels are surrounded 
by a row of extra cupels so that the temperature may be kept as 
uniform as possible for all the assays. The cupels are withdrawn 
while the buttons are still fluid. With a two to one ratio of silver to 
gold, and with copper present, there is no danger of sprouting. 

The buttons are removed from the cupels by means of pliers and 


132 


carefully cleaned from all adhering bone-ash. They are then placed 
on a special anvil and flattened by a middle and two end blows with 
a heavy polished hammer. They are then annealed at a redheat and 
passed twice through the rolls which are adjusted each time so that 
after the second passage they are about 23’’ long by 4’’ wide, and 
about as thick as an ordinary visiting card. It is important that the 
fillets be all of the same size and thickness with smooth edges. They 
are then re-annealed and each one is numbered on one end with small 
steel dies to correspond with the number of the assay, and rolled up 
into ‘“cornets” or spirals between the finger and thumb, with the 
number outside. It is important that an even space be left between 
all turns of the spiral in order that the acid shall have easy access 
to all parts of the gold. 


The cornets are parted in platinum thimbles which are supported — 


in a platinum basket, and the whole thing is placed in a platinum 
vessel containing boiling nitric acid of 32° B. (Sp. Gr. 1.28). They 
are boiled for 10 minutes and then transferred to another vessel 
containing acid of the same strength and boiled 10 minutes longer. 
The basket and contents is then washed by dipping vertically in and 
out in three changes of distilled water, drained, dried, and annealed 
usually in the muffle. 

When cold the cornets are ready to weigh. The gold should be 
entirely in one piece, and the original numbers easily discernable on 
the parted cornets. The proofs are weighed first and the corrections 
applied to other cornets. The proofs always show a slight gain in 
weight. The correction thus determined is termed the be 
and is really the algebraic sum of all the gains and losses. 

When more than 14 cornets are parted at one time the lot is given 
a preliminary 3 minute treatment in an extra lot of acid followed by 
the two regular 10 minute boilings. 

The purpose of the copper which is added to the assays is to render 
the button tough and permit of its being rolled out into a smooth 
edged fillet. Without the copper, the fillet is apt to crack in rolling, 
or to come through with a ragged edge which might give rise to a loss 
in parting. The action of copper in this case is probably due to its 
effect in aiding in the removal of the last of the lead in cupelling.1 
The time required for cupellation is approximately 12 minutes. 

1 Rose Trans. Inst. Min. & Met. 14 pp. 545. 


ES 








CHAPTER XI. 


THE ASSAY OF SOLUTIONS. 


A large variety of methods for the assay of gold and silver bearing 
solutions have been published in the technical press, and quite a 
number of these have been adopted by assayers. These methods 
may be classified as follows:— 

1. Methods involving evaporation in lead trays with subsequent 
cupellation, or scorification and cupellation of the tray and contents. 

2. Methods involving evaporation with litharge and other fluxes 
followed by a crucible fusion and cupellation. 

3. Methods in which the precious metals are precipitated and either 
cupelled directly or first fused or scorified and cupelled. 

4. Electrolytic methods in which the precious metals are deposited 
directly on cathodes of lead foil, which are later wrapped up with 
the deposit and cupelled. 

5. Colorimetric methods (for gold only) all of which depend upon 
obtaining the “‘purple of Cassius’’ color which may be compared with 
proper standards. 


Evaporation in Lead Tray. This method is a good one on rich, 


neutral solutions containing only salts of the precious metals. A 


tray of suitable size is made by turning up the edges of a piece of lead 
foil. If many of these assays are to be made it is well to have a wooden 
block as a form on which the trays may be shaped. A tray 2’’ X 
2’’ x 2#’’ deep is about right to hold 1 assay-ton of solution. 

Having made a tray which will not leak, the solution is added and 
carefully evaporated to prevent spattering. The tray is then folded 
into a compact mass and dropped into a hot cupel. 

Among the disadvantages of the method are the following: It 
does not permit of the use of a large quantity of solution, and there- 
fore is suited only to rich solutions. If the solutions are acid they 
will corrode the tray, and if they contain salts other than those of 
gold and silver these will interfere with cupellation. As both AuCl, 
and KAu(CN), are volatile at moderate temperatures, many assayers 


- do not consider the method a reliable one for solutions of these salts. 


Evaporation with Litharge. (First Method.) A measured 
quantity of the solution is placed in a porcelain evaporating dish and 


134 


from 30 to 60 grams of litharge is sprinkled over the surface. The 
mixture is allowed to evaporate at a gentle heat to prevent both 
spitting and baking of the residue. When dry the residue is scraped 
out, mixed with suitable fluxes, transferred to a crucible and fused in 
the ordinary manner. The last portions remaining on the dish may 
be removed by means of a small piece of filter paper slightly moistened 
which is afterwards added to the charge. 

Some assayers add a little fine silica and charcoal with ihe litharge. 
The soluble constituents of a crucible charge, soda and borax, should 
not be added to the solution as they form a hard cake which is difficult 
to remove from the dish. The most important point in the process 


_ is the proper manipulation of the temperature. If this is right there 


will be no spattering and the dry residue will come away practically 
clean from the dish after prying It up with the point of a spatula. 


Evaporation with Litharge. (Second Method.) A measured 
amount of solution is evaporated in a porcelain or enameled iron dish 
to a small volume, so that when the litharge and silica of a crucible 
charge are added, they will absorb practically all of the liquid forming 
a thick paste. The heating is continued and the material is stirred 
constantly to keep the residue granular, and to prevent it from stick- 
ing to the dish. When dry the residue is cleaned out with the aid of 
a spatula and a mixture of soda, borax and are argols placed with it 
in a crucible which is heated to quiet fusion, poured and treated in 
the usual way. . 

This method requires more manipulation than the first one, and 
the only advantage is in a possible hastening of the process. 

If any residue sticks to the dish it may be removed by rubbaiee it 
with a little fine silica on a piece of filter paper, the whole being after- 
wards added to the charge. 

A modification of the foregoing evaporation methods consists in 
evaporating to a small volume without the addition of any reagents, 
and then transferring the concentrated solution to a small dish of 
very thin glass (Hoffmeister’s dish). The solution is evaporated 
to dryness either with or without litharge, and the dish and contents ~ 
broken up directly into a crucible containing the usual fluxes. The 
assay is finished in the usual manner. The advantage of this method 


lies in the fact that there is no chance for loss of residue by not properly _ 


cleaning the dish, as the dish and all are fused. 

The evaporation method while somewhat long, is the most reliable 
and accurate one known, and is the standard with which all other 
methods are compared. By arranging to allow the evaporation to 








135 


run over night, the samples taken one night may be assayed and re- 
ported early next morning. The method is adapted to the treatment 
of any quantity of solution and of almost any character. If the solu- 
tion contains much sulphuric acid, the litharge may be converted 
into lead sulphate which is unsuited either to act as a flux or to pro- 
vide lead for a collecting agent. A fusion made on such a substance 
using a carbonaceous reducing agent, will give either no button at 
all, or a button of matte. The reaction between lead sulphate and 
carbon is as follows: ) 
PbSO; + 2C = PbS + 2CO, 

If the solution is one of AuCl; a little charcoal should be added dur- 
ing the evaporation to insure the reduction and precipitation of the 
gold, asin this way we avoid the danger of loss gold by volatilization 
as the chloride. The gold being precipitated on the charcoal is in 
the best possible position to be alloyed with the lead which will be 
reduced by the carbon. 

Precipitation by Zinc and Lead Acetate. The Chiddey 
Method. (For Cyanide Solutions.) This method which was first 
described by Alfred Chiddey ' is suitable for both gold and silver and 
is used almost exclusively in this country for the assay of cyanide 
solutions. It works equally well on strong or weak, foul or pure 
solutions, and almost any quantity may be taken. Many changes 
of detail have been suggested and innumerable modifications of the 
original process are found described in the technical press. The 
following method has been found satisfactory: 

Take from one to twenty assay-tons of solution in a beaker or 
evaporating dish and heat. Add 10 or 20 c.c. of a 10 per cent solution 
of lead acetate containing 40 c.c. of acetic acid per liter. Then add 
one or two grams of fine zine shavings rolled lightly into a ball. The 
gold, silver and lead will immediately commence to precipitate on 
the zinc. At first the solution may become cloudy but will soon clear 
as more of the lead is precipitated. Heat, but not to boiling until 
the lead is well precipitated. This usually takes about twenty or 
twenty-five minutes. Then add slowly (about 5 c.c. at a time), 
20 c.c. hydrochloric acid (1.12 sp. gr.), to dissolve the excess zinc. 
Continue heating until effervescence stops. It is often found that 
action ceases while there is still some undissolved zine remaining. 
This is entirely covered and thus protected from the acid by the spongy 
lead. To be sure that all the zinc is dissolved feel of the sponge with 
a stirring rod and drop a little hydrochloric acid from a pipette di- 
rectly on it. 

1H. & M. J. 75, p. 478, March 1903. 


136 


As soon as the zine is dissolved decant off the solution and wash 


the sponge two or three times with tap water. Next, moisten the | 


fingers and press the sponge, which should be all in one piece, into a 
compact mass. Dry by squeezing between pieces of soft filter paper 
or by placing on a piece of lead foil and rolling with a piece of large 
glass tubing. Finally roll into a ball with lead foil, puncture to allow 
for steam escape, add silver for parting, and place in a hot cupel. 

As soon as the zine is dissolved the assay should be removed from 
the heat, and the sponge removed. If this is not done the lead will 
start to dissolve and the sponge will soon break up. Washing by 


decantation and manipulation with the fingers may appear crude, - 


but after a little practice the operator becomes so proficient that 
there is practically no chance of loosing any of the lead. 

If any considerable amount of water is left the assay will split in 
the cupel. To avoid this danger some assayers dry the assays on the 
steam table before cupelling. Any zinc left will also probably cause 
spitting. Chiddey recommends placing a piece of dry pine wood in 
the mouth of the muffle immediately after changing the cupels, 
probably with the idea that this aids to prevent spitting when some 
zinc has been left undissolved. When working with small quantities 
of solutions it is best to add water occasionally to maintain a volume 
of at least 100-150 ¢.c. The secret of keeping the lead from breaking 
up is not to allow the solution to come to a boil at any stage of the 
procedure. ? 

Zine dust is used by many chemists in place of zine shavings, a 
small amount being added on the end of a spatula. Many chemists 
agree that 4 gram is sufficient. 

William H. Barton‘ suggests the addition of a small piece of alum- 
inum foil dropped into the solution after the hydrochloric acid is 
added, to prevent the dissolving of the lead and the consequent 
breaking up of the sponge by the hydrochloric acid after the zinc is 
all dissolved. 

T. P. Holt * recommends the substitution of a square of aluminum 
foil for the zinc. The lead sponge is removed from the aluminum 
with a rubber-tipped stirring rod. Care must be taken to use a 
sufficiently thick sheet of aluminum (1/16’’ does well), to prevent 
small pieces becoming detached. These would remain with the lead 
sponge and might cause the cupels to spit. 

Precipitation as Sulphide.’ Acidify five or ten assay tons 

1! Western Chemist and Metallurgist, 4 p. 67, Feb. 1908. 


* Mining & Scientific Press, 100 p. 863, June 1910. 
3 Henry Watson, E. & M. J. 66 p. 753, Dec. 1898. 


— ee re ee a ee ee Se 








137 


of solution with HCl and heat to boiling. While boiling add a solu- 
tion containing two grams of lead acetate and pass in a current of 
hydrogen sulphide until all the lead is precipitated. Allow to cool 
somewhat still passing in H2S, then filter and dry. Collect the gold 
and silver with lead either by a crucible fusion or scorification assay. 
The method is said to be quick, accurate and economical. 


Precipitation by Cement Copper.’ To 8 assay tons of the solu- 
tion add a few cubic centimeters of sulphuric acid, and one gram of 
finely divided cement copper. Hear to boiling and boil 10 minutes. 
Filter through a strong 7 inch paper and place on the drained filter 
one-third of a crucible charge of mixed flux. Place the filter in a 
crucible containing another third of a charge of flux, and cover with 
the final third. Fuse and cupel as usual. The filter itself furnishes 
the reducing agent for the assay. If cement copper is not available, 
a solution of copper sulphate may be added together with a small 
piece of aluminum foil. Boil until all the copper is precipitated and 
add the remaining aluminum foil to the fusion. This modification 
takes more time than the first. 


Precipitation by Silver Nitrate (For Gold in Cyanide Solu- 
tions). Add an excess of silver nitrate solution which will cause the 
gold and silver to precipitate as an auric-argentic-cyanide. Allow 
the precipitate to settle, filter through a thin paper, and wash several 
times. Dry the filter and either scorify with test lead or fuse in a 
crucible with litharge and the regular fluxes. The method gives fairly 
good results in solutions not too low in gold. With solutions very 
low in gold the precipitation of the gold is not perfect. 


Precipitation by a Copper Salt’ (For Cyanide Solutions Only). 
Add to one liter of solution in a two liter flask 25 c.c. of a 10 per cent 
solution of copper sulphate, then add 5 to 7 ¢c.c. of concentrated hydro- 
chloric acid and lastly 10 to 20 ¢c.c. of a 10 per cent solution of sodium 
sulphite. Shake vigorously for at least two minutes then filter, 
dry, and fuse the filter and precipitate in the usual way. With weak 
solutions it is best to bring up the strength by the addition of cyanide 
_ before adding the copper salt. The gold and silver are carried down 
by the precipitate of cuprous cyanide formed. Assays may be 
completed in three hours and the results are said to be good on both 
low and high grade solutions. 

Albert Arents, T. A. I. M. E. 34 p. 184. 


1 
2 Andrew F. Cross, Jl. Chem. Met. and Min. Soe. of So. Af. 1 p. 28, and 3 p. lL. 
3 A. Whitby, Proc. Chem. Met. and Min. Soe. of So. Af. 3 p. 6. 


138 


The Electrolytic Assay of Cyanide Solutions. The following 
method is abstracted from the Journal of the Chemical, Metallurgical 
and Mining Society of South Africa ' which describes the method and 
installation used at the Kleinfontein Group Central Administration 
Assay Offices. 

Ten-assay-ton samples of the solution to be assayed are placed in 
No. 3 beakers, which are held in‘a frame, and electrolyzed using a 
current of 0.1 ampere. The anodés used consist of ordinary 5/16 
inch arc lamp carbons which are held in position in the center of each 
beaker by suitable clamps. They are arranged so that they may be 
lifted out of the solution when no current is passing. The cathodes 
are made from strips of ordinary assay lead foil 23’’ x 9’’ with the 
lower edge coarsely serrated to allow for circulation of the solution. 
To connect with the battery a }’’ 
of the foil, and turned upward to make a terminal. The two ends of 
the lead are brought together and connected by folding the edges, 
making a cylinder about 3 inches in diameter. 

The time required for the complete deposition of the gold is four 
hours, after which the carbons are removed, the lead cathodes dis- 
connected and dried on a hot plate. When dry, they are folded into 
a compact mass and cupelled. 

With weak solutions a small quantity of cyanide should be added 
in order to decrease the resistance and thus accelerate the deposition 
of the precious metals. The author reports no difficulty in obtaining 
a complete and adherent deposit of the gold, which separates as a 
bright yellow deposit. 

This of course was the only metal worked for on the Rand, but 
there seems to be no reason why silver as well as gold can not be de- 
termined by this method. 

The principal advantage of the method lies in the small amount of 
actual personal attention required. The method works as well 
for a 20 A. T. sample as for one of 10 A. T. The time required for 
the deposition of the gold is somewhat longer than for some of the 
precipitation methods and this appears to be the principal disad- 
vantage of the process. 


Colorimetric Methods. (For Gold only.) Several attempts 
have been made to adapt the ‘“‘Purple of Cassius”’ test to the estimation 
of gold in chlorine and cyanide solutions. So far as the author is 
aware, none of the methods have been adopted as practical assay 
laboratory methods in this country. They were used for a time in 

1 Vol. 12 p. 90. C. Crichton. 


strip is all but severed from one end | 





a ee 


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=< = 


a 
Peni 
er aa ae ee hee vi) 










aaa 139 
one for two South African plants, but have never come into great 
favo r. The two most promising methods were described by Henry 
R. Cassel (E. & M. J. 76 p. 661) and James Moir (Proc. Chem. 
. and Min. Soc. of So. Af. 4 p. 298) and to those original articles 


interested reader is referred. 


an. ha ay pa A > 
¥ sa Ne A ba > ‘ 


we 7 


CHAPTER XII. 


THE LEAD ASSAY. 


The fire assay for lead consists of a reducing fusion with iron, 
fluxes, and some carbonaceous reducing agent, and is conducted much 
as is the iron nail assay for gold and silver ores except of course no lith- 
arge or other lead bearing fluxis added. The object of the fusion is to 
reduce and collect all of the lead in a button free from other 2lements. 


Lead Ores. Lead ores are classified by metallurgists as oxidized 
or sulphide ores, also as pure or impure ores. The oxidized ores 
contain the lead principally in the form of carbonate, occasionally 
as sulphate and rarely as oxide or in combination with phosphorous, 
molybdenum, vanadium, chromium, etc. The corresponding lead 
minerals are cerussite, PbCO; (77.6% Pb), anglesite PbSO, (68.3% 
Pb), minium Pb3O, (90.6% Pb), pyromorphite PbsCl (POx)3 (75.6% 
Pb), vanadinite 3Pb3(VOz)2 PbCl, (72.4% Pb) and wulfenite PhMoO, 
(56.5% Pb). The most important sulphide lead minerals are galena 
PbS (86.6% Pb) jamesonite PbeSb.S, (50.8% Pb) and bournonite 
PbCuSbS; (42.5% Pb). The principal associated minerals are 
argentite, pyrite, chalcopyrite, sphalerite, stibnite, quartz, calcite 
and dolomite, as well as the oxidation compounds of the above sul- 
phides. Impure ores, from the assayers point of view are those con- 
taining more or less arsenic, antimony, bismuth, copper, zine, and 
other rarer metals which interfere with the lead assay. 

Besides ores, the assayer may have brought to him various furnace. 
products such as litharge, slag, matte, flue dust and cupel bottoms. 

The fire assay for lead is not always as accurate as a carefully made 
wet determination but it is so simple, inexpensive and rapid that for 
a long time it served to govern the purchase and sale of all lead ores. 
Today it is still largely used by the smelters and others for the assay 
of pure ores, although for ores containing such base metal impuritie: 
as antimony, copper, zinc, etc. the wet method is usually preferred. 

The results of the fire assay may be either lower or higher than the 
actual lead content, depending on the nature and quantity of the 
other minerals present in the ore. Pure ores give low results owing 
to losses of lead by volatilization and slagging. Both the sulphide 
and oxide of lead are volatile at moderate temperatures and for this 
reason great care must be taken to keep the temperature as low as 


141 


possible consistent with a proper decomposition of the ore, and of the 
lead compounds which are formed during the fusion. On the other 
hand lead compounds, particularly the oxide, tend to pass into the 
slag which tendency is increased by the presence of zinc, and to some 
extent by arsenic and antimony. Impure ores containing arsenic, 
antimony, bismuth and copper, usually give high results as these 
metals are partly or wholly reduced and pass into the lead button. 


Quantity of Ore and Reagents Used. The amount of ore 
used is generally 10 grams, occasionally 5 grams. With low grade 
ores 20, 25, or more grams may be used. ‘The reagents used are the 
alakli carbonates, borax-glass, some reducing -agent usually argols 
or flour, and occasionally sulphur. Iron in some form is always used. 
It may be in the form of nails or spikes, coiled wire, or the crucible 
itself may be of iron, and in this case will be used over and over again 
until worn out. A very satisfactory way of introducing iron is to 
use a rail or boat spike 25 or 3 inches long, and about 0.5 inches 
through. In this assay it is customary to use a mixture of sodium 
and potassium carbonates as the mixture fuses at a lower temperature 
than either one alone. The alkali carbonates act as fluxes for the 
silica, and serve to give a basic slag which is necessary in this assay. 
Usually from two to three times as much alkaline carbonate as ore 
is taken. Borax-glass acts as a flux for the metallic oxides for lime- 
stone and the other alkaline earths. From one-half to twice as much 
borax-glass as ore is used. An excess of reducing agent 1s always used 
to insure keeping the necessary strongly reducing condition. Sul- 
phur is only occasionally used and then when assaying an oxidized 
ore containing copper. 

In the lead assay it is customary to use a mixed flux called a “‘lead 
flux.’ This may be bought already prepared or may be made up in 
the laboratory. Many different formulas are given among which are 
the following: 


L 2 o. 
Sodium-bicarbonate 12 parts 4 parts 6.5 parts. 
Potassium carbonate 1Dtae ria Aine 
Borax-glass ny ime — 26D ies 
Borax powdered = Chae _ 
Flour Zsa Liven A te 


No. 1 and 2 are found in use in the Coeur d’ Alene lead district 
where the fire assay for lead has been brought to the highest degree 
of perfection. No. 1 is better for ores having a basic gangue, No. 2 
for siliceous ores. No. 3 is perhaps the best of all for general use. 


Fa 


142 


About 380 grams of flux are intimately mixed with 10 grams of ore, 
a nail is inserted and a cover of 8 or 10 grams more of flux is added. 
Very few assayers use a cover of salt in the lead assay on account of 
the danger of the loss of lead as chloride. 

The fusion should always be made in a muffle furnace owing to the 
better control of temperature available. In fact the secret of the 
successful fire assay for lead is largely in the proper manipulation 
and control of the temperature throughout the process: 


At the start the temperature should be low, sufficient only to barely _ 


melt the charge. This is necessary owing to the fact that in the early 
part of the assay the charge is in active motion and particles of the 
various lead compounds are continually being brought to the surface, 
where if the temperature were high they would suffer an appreciable 
loss by volatilization. When the charge has finished boiling and most 
of the lead is reduced and collected in the bottom of the crucible 
there is less danger of a loss by volatilization, owing first to the fact 
that lead itself is not so readily volatile as are some of its compounds 
and second to the difficulty of migration of the molecules through the 
heavy layer of reducing slag which covers the lead. 

After the boiling has entirely ceased the temperature is raised gradu- 
ally to decompose the lead compounds which still remain in the slag. 
These are principally the silicate and the double sulphide of lead and 
sodium or potassium and require a bright yellow heat for their com- 
plete decomposition. The fusion period is finished when the nails 
can be removed free from shots of lead. Sulphide ores require a much 
longer fusion than oxides owing to the fact that their decomposition 
is effected principally by iron, and therefore time must be allowed for 
every particle of the charge to come into contact with the iron. 
Oxide ores, on the other hand, are decomposed by the carbon of the 
charge and as this is uniformly distributed a much shorter time will 
suffice. Sulphide ores will require from one to one and one half 
hours of fusion, oxide ores from three quarters of an hour to an hour. 

Influence of Other Metals on Lead Assay. Silver. Practically 
all of the silver in an ore is reduced and passes into the lead button. 
If present in sufficiently large quantities a correction for it may be 
made, i.e., 291.66 oz. per ton equals one per cent. 

Gold. This metal is also reduced and passes into the lead button, 
but it is usually present in such small quantities that it may be dis- 
regarded. 

Arsenic. Arsenic is occasionally found in lead ores usually in the 
form of arsenical iron pyrite. During the assay part of the arsenic 


is volatilized as metal or as arsenic sulphide but the larger part re- 


* 


eee ee _ 





143 


mains in the crucible. Here it usually enters into combination with 
the iron forming a speiss. After pouring it will be found as a hard 
white button on top of the lead from which it may be removed by 
hammering. Little if any arsenic enters the lead button. Under 
certain conditions, 1.e., a long fusion at a low temperature with high 
soda excess, the formation of a speiss may be prevented. 


Antimony. This metal is frequently found associated with lead, 
usually however only in small amounts. In the assay with iron, 
antimony is reduced and passes into the lead button. Buttons con- 
taining antimony are harder and whiter than those from pure lead 
ores and when they contain much antimony are brittle, breaking with 
a bright crystalline fracture. 

If much antimony is present (over one-half as much as the lead) 
an antimony speiss will be found lying on top of the button. 
Bismuth. This metal is rarely found associated with lead ores, but 
if present will be reduced and pass into the lead buttons. 


Copper. Copper is often found in lead ores in the form of chalco- 
pyrite, chalcocite, and oxidized copper compounds. If the ore is 
fully oxidized and a high temperature is employed most of the copper 
will pass into the lead button. If the ore contains much pyrite or 
sulphur in other forms most of the copper will remain as a sulphide 
and be dissolved in the alkaline slag. A button containing copper 
will be hard and tough and may show a reddish tinge. 

Iron. This metal is often present in lead ores usually in the form 
of iron pyrite. It goes into the slag forming either a silicate or a 
double sulphide of iron with sodium or potassium. The lead button 
is practically free from iron. 

Zinc. Zine is often found associated with lead in ores usually 
in the form of the sulphide. During the assay part of the zinc is 
volatilized and part remains in the slag. Zine sulphide is only de- 
composed by iron at a very high temperature so that only a very 
small amount of zinc passes into the lead button. Zine sulphide is 
practically infusible, so that if present in too great an amount, may 
make the slag thick and pasty, and thus interfere with the separation 
of the lead. 


Procedure. Assay ores in duplicate using 10 grams of ore and 40 
grams of prepared lead flux. Use a 12 or 15 gram muffle crucible. 
Weigh out first 30 grams of lead flux, place the ore on top of this and 
mix thoroughly with the spatula. Insert a spike or nails point down- 
ward and finally cover with 10 grams more of lead flux. Have the 


144 


muffle only moderately red so that it will take at least 30 minutes 
from the time the charges are put in until they are boiled down. 
Close the door to the muffle as soon as the crucibles are in and after 


the charges are melted place two crucibles part full of soft coal in — 


the mouth of the muffle just inside of the door, which should be kept 
as tightly closed as possible. Raise the temperature gradually to a 


bright yellow and continue at this temperature until the nails can be 


removed free from lead. | 

Finally take the crucibles from the muffle using a pair of muffle 
crucible tongs and without setting them down quickly remove the 
nails with a large pair of steel forceps, tapping against the side of the 
crucible and washing the nails in the slag to remove all adhering lead 
globules. Pour the fusion into a mold and when cool separate the 
lead from the slag and hammer clean. Weigh to centigrams and 
report the results in percentage. Duplicates should be checked 
within 0.5 per cent. 

The slag should be black and glassy. If dull, more borax-glass 


should be added. It should pour well from the crucible and immedi- ~ 


ately after pouring, the crucible should be examined for shots of lead. 
If these are found it is usually an indication of too low a temperature 
at pouring. 


Notes. 1. If the ore is an oxide and contains copper add a gram or two of finely 
pulverized sulphur to the charge to prevent the copper from entering the button. 

2. The soft coal is added to insure reducing conditions in the muffle and it may 
be renewed if necessary. When a muffle is used solely for fusion purposes the hole 
in the back is stopped up, thus preventing the entrance of so much air. 

3. The removal of nails and pouring must be done without a moments delay as 
the charges are small and cool rapidly. 

4. If the ore contains much silver the button should be cupelled and the weight of 
silver found deducted. 

5. The lead should be soft and malleable and a fresh cut surface should have the 
bluish gray color of pure lead. The button should be capable of being hammered 
out into a thin sheet without breaking or cracking. A button that is bright, brittle 
and brilliantly white in the fracture indicates the presence of arsenic or antimony. 

6. If there is doubt regarding the purity of the lead button it may be tested by 
cupellation. The only metals except lead likely to be present are gold, silver, anti- 
ee copper and possibly zinc; each of which gives characteristic indications in 
cupelling. 

7. Crucibles may be used a number of times as they are but little corroded but 
those used previously for gold and silver assays must nol be used for this assay as the 
slag left in them contains lead. It is well to use a special size of crucible for the lead 
assay in order to prevent errors from mixing crucibles. 


Assay of Slags, Furnace Products and Low Grade Ores or 
Tailings. In the assay of low grade materials such as slags and 
tailings a larger quantity of ore should be used and a different mixture 
of fluxes. The slag should be between a singulo and a sub-silicate 
and part of the iron may be added in the form of filings. On account 
of the size of the charge it is well to add a number of nails, as this 
will lessen the time necessary for complete reduction. 





The following charges have been found satisfactory : 


145 


Limestone (3-2% Pb) Slag Slag 
Ore 25 gm. Slag 25 gm. Slag 100 gm. 
NasCO; 25 ay NaeCO3 20 y NaeCO; 50 
K.CO; 20 os K.COs; 20 a K.CO; mr 
Borax-glass 20 ‘“ Borax-glass 10 ‘“ Borax-glass 10 “ 
Flour Pose Flour Mg Se Nelielnte LOP os 
Nails ree Nails Det pee Nets 5 els 
(20 penny) (20 penny) (20 penny) 


20 gram crucible 


20 gram crucible - 


30 gram crucib!e 





Allow some time at a high temperature to allow opportunity for 
all of the slag to come in contact with the iron. 

Corrected Lead Assay. To recover any lead which may have 
been left in the slag the following procedure is recommended. Save 
all the slag and remelt in the original crucible with the spikes or nails 
formerly used. If the first slag was quite glassy and viscous in 
pouring, add from 5 to 15 grams more sodium carbonate. Heat to 
redness and drop into each crucible a lump of about 5 grams of 
potassium cyanide. Close the door to the muffle and heat to a bright 
yellow and pour as soon as quiet. Add the weight of any small 
- button found to the lead from the original fusion. 


Chemical Reactions of the Lead Assay. 


With an ore containing PbCOs;, PbSO., PbS, SiO. and CaCO; the 
following reactions may occur :— 
PbCO; = PbO + CO, 
2PbO=- C = 2Pb + CO, 

PbO + S810, = Pb SiO; *(Begins at 625° C.) 
PbSO, + 2C = PbS + 2CO, (Begins at a dark red heat.) 
7 PbS + 4K.CO; = 4Pb + 3 (K2PbS,) + K,SO, + 4CO2 

(Begins at a red heat.) 

If carbon were not present some oxide and sulphate would probably 
-remain to react as follows :— 

PbS + 2PbO = 3Pb + SO, (Begins at 720° C.) 
PbS + PbSO, = 2 Pb + 250, (Begins at 670° C.) 
2PbSO, + SiO2 = Phe SiO, + 2802 + O2 (High heat.) 

Toward the end as the heat is raised to a bright red and above, the 
reactions with iron become of importance, particularly the following :— 
PbS + Fe = Pb + FeS 
PbSi0; + Fe = Pb + FeSiO; 


(Begins at 200° C.) 
, (Begins at 550° C.) 


(Requires a bright yellow: heat for 
completion. ) 
(Requires a bright eo heat for 
completion. ) 


K»PbS + Fe = Pb + KeFeS: 





INDEX 


Acid slags, 88. 
Annealing parted gold, 69, 70. 
Antimony, effect in iron nail assay, 108. 
effect in lead assay, 143. 
effect in scorification, 77, 78. 
Antimonial ores, crucible assay of, 116. 
Argols, 4, 5. 
Arsenic, effect in iron nail assay, 108. 
effect in lead assay, 142. 
effect in scorification, 77, 78. 
Assay—ton weights, 48. 


Balance, alignment of knife edges, test- 
ing, 46. 
arms, equality, testing, 46. 
button, 38. 
construction of, 38. 
directions for use of, 40, 41. 
equilibrium, testing, 44. 
flux, 37. 
multiple rider attachment, 47. 
pulp, 37. 
sensibility of, 45, 46. 
stability of, 45. 
testing an assay balance, 44-46. 
theory of, 38-40. 
time of oscillation, 45. 
Basic ores, assay of, 92. 
calculation of charge for, 92. 
slags for, 92, 93. 
Basic slags, 88. 
Bone-ash, 51, 52. 
best size for cupels, 52. 
temperature of burning, influence of, 
51. 
Bone-ash cupel, assay of, 118. 
Borax, 2. 
effect of in slags, 86-88, 98. 
use of in assaying, 86-88, 98, 102. 
Borax-glass, 2. 
Bismuth, effect in lead assay, 143. 


Bullion, copper, assay of, 123-128. 
doré, 128-130. 
gold, 130-133. 
lead, 123. 
nomenclature, 119. 
sampling of, 119-123. 
silver, 128-130. 


Capsules, parting, 69. 
Character of sample, determination of, 
85. 
Charcoal, 5. 
Chiddey method for assay of cyanide 
solution, 135, 136. 
Class 1 ores, assay of 89-101. 
Class 2 ores, assay of, 97-110. 
Class 3 ores, assay of, 110. 
Classification of ores, 84. 
Classification of silicates, 86. 
“Cleaning the slag” in scorification, 82. 
Coal furnaces, 12. 
firing of, 12. 
Cobalt, effect in scorification, 78. 
Coke furnaces, 13. 
Colorimetric method of assay, 138. 
Combination assay of copper bullion, 
123, 124, 126, 127. 
Copper bullion, assay of, 123-128. 
sampling, 121-123. 
Copper, crucible method for ores high 
in, 114, 116. 
effect of in cupellation, 62, 63. 
effect of in iron nail assay, 108. 
effect in lead assay, 143. 
effect in scorification, 77, 78. 
matte, assay of, 80. 
Corrected assays, gold and silver, 117, 
118. 
lead, 145. 
Cover, the, 98. 
Cryolite, 7. 


148 


Crucible assay, 83. 
copper bullion, 123, 125, 126. 
procedure for, 100, 101, 103-107. 
theory of, 83, 108-111. 
Crucible furnaces, 9. 
Crucible slags, ‘properties of, 85. 
Crucibles, 18, 19. 
capacity of different sizes, 19. 
desirable properties of, 18. 
size of for various charges, i11. 
Cupels, 51, 52. 
assay of, 118, 121. 
cracking of, 52, 53. 
effect of shape of, 54. 
instructions for making, 52, 53. 
machines for making, 53. 
magnesia, 66. 
Portland cement, 66. 
size of, 53, 54. 
testing, 53, 65, 66. 
Cupellation, 51, 54-57. 
appearance of buttons containing 
platinum, 58, 65. 
correct temperature for, 55. 
‘freezing’ of button in, 56, 58. 
indications of metals present, 64. 
instructions for, 57-59. 
loss of gold in, 61. 
loss of silver in, 59-61. 
regulation of temperature during, 57, 
58. 
retention of base metals in beads 
from, 66. 
“spitting” during, 53. 
“sprouting”’ of silver after, 56. 
Cupellation losses, 59-64. 
influence of copper on, 62, 63. 
influence of impurities, 62. 
influence of quantity of lead, 60. 
influence of tellurium, 113. 
influence of temperature, 59-61. 
Cyanide solutions, assay of, 133-139. 


Doré bullion, assay of, 128-130. 


Electrolytic assay of cyanide solutions, 
138. 

Fire brick, directions for laying, 16. 

Fire brick lining vs. tiles for lining, 11. 


Flasks, parting, 71. 
Flour, 5. 

Fluidity of slags, 88. 
Fluorspar, 7. 


Fluxes and reagents, 1-7. 
Fluxing, principles of, 84. 
Fuel, advantages of gas and oil over 
solid, LW 
for assay furnaces, 10, 11. 


Furnace repairs, 15, 16.' 

Furnaces, 9-15. 
directions for firing, 12. 

Fusion products, 7, 8. 


Gasolene furnaces, 13, 14. 

Gold, ores containing coarse particles, 
assay of, 35, 36. 

Gold bullion, assay of, 130-133. 

Gold, weighing, 69, 70. 

Granulated lead, assay of, 79. 


Heat of formation of metallic oxides, 
74, Pia: 


Ignition temperature of metallic sul- 
phides, 74. 

Inquartation, 70, 71. 

Iron, 5. 
effect in lead assay, 143. 
effect in scorification, 77, 78. 
reducing action of, 5. 

Iron assay, 103, 107-109. 


Jones sampler, 26. 


Lead, 4. 
fire assay for, 140-145. 
granulated, assay of, 79. 
granulated to make, 4. 
ores, 140. 

Lead bullion, assays of, 58, 59, 123. 
sampling, 121. 

Litharge, 3, 4. 
assay of, 99, 100. 
corrosive action of, 75. 
Lodge’s rule for, 106. 
solubility of metallic oxides in, 73, 


74. 
use in the scorification assay, 75. 


Lodge’s rule for the use of litharge, 106. 





Magnesia cupel, assay of, 118. 
Manganese, effect in scorification, 78. 
Matte, 8. 
copper, assay of, 80. 
Mercury-sulphuric acid combination 
method for copper bullion, 123, 
124, 126, 127. 
Metallic assay, 35, 36. 
Metallic oxides, heat of formation of. 
74, 113. 
solubility in litharge, 73, 74. 
Metallic sulphides, ignition tempera- 
ture, of 74. 
Minerals, oxidizing power of, 96. 
reducing power of, 96. 
Moisture sample, 22. 
Muffles, 17. 
care of, 67. 
directions for replacing, 16. 
method of supporting, 12. 
spilling in, 67. 
Muffle furnaces, 9, 11, 12. 
Multiple rider attachment for balances, 
47. 


Neutral ores, 84. 
Nickel, difficulty of eliminating in 
scorification, 74. 
effect of in iron nail assay, 108. 
effect in scorification, 74, 79. 
Niter, 5, 6. 
oxidizing power of, 105. 
oxidizing reactions, 5, 6. 
Niter assay, 102-107. 
objections to, 107. 
preliminary fusion, 105-107. 
regular fusion, 105-107. 


Nitric acid combination method for 
copper bullion, 124, 127. 


Ores, classification of, 84. 
determining reducing power of, 103, 
104. 
estimating reducing power of, 104, 
105. 
Oxides metallic, heat of formation of, 
7a) 118. | 
solubility in litharge, 73, 74. 


149 


Oxidizing power, definition, 94. 
of minerals, 96. 
of niter, 97, 105. 
of niter, determining, 105. 
power of ore, to find, 111. 
ores having, 84. 
of red lead, 97. 

Oxidizing reactions, 96, 97. 


Parting, 68. 

acid for, 68, 69, 133. 

capsules, 68. 

flasks, 71. 

preparing beads for, 71. 

procedure, 69-72, 133. 

ratio of silver to gold necessary, 68, 

70. 

Portland cement cupel, assay of, 118. 
Potassium carbonate, 3. : 
Potassium cyanide, 6. 


Reactions, during iron assay, 107, 108. 
lead assay, 145. 
oxidizing, 96, 97. 
reducing, 94, 95. 
Reagents, 1-7. 
testing, 98, 99: 
Reducing agents, 4, 5. 
Reducing power, definition, 94. 
of minerals, 97. 
of ores, determination of, 103, 104. 
of ores, estimation of, 104, 105. 
ores having, 84. 
of reagents, 94. 
of reagents, determination of, 99. 
Reducing reactions, 94—96. 
Rescorifying buttons, 78, 80, 124. 
Riders, 47, 50. 
Riffle cutter, 26. 
Roasting method, 103, 109, 110. 


Salt, action of, 98. 
objections to, 98. 
Sample, definitions, 21. 
finishing, 34, 35. 
moisture, 34, 35. 
Samples, labelling, 22. 
Sampling, copper bullion, 121-123. 
duplicate, 34. 


150 


Sampling, gold bullion, 130. 
hand and machine compared, 27. 
lead bullion, 121. 
Sampling, machine, 26, 27. 
operations, 22~27. 
ores containing malleable minerals, 
35, 36. 
Brunton’s formula for, 32. 
Reed’s formula for, 31. 
Richard’s rule for, 29, 
tables showing weights to be taken, 
30-33. 
theory of, 28-32, 
Scorification assay, 74. 
charges for different ores, 81. 
chemical reactions in, 77, 
of copper bullion, 123-125, 
of copper matte, 80. 
effect of various metals, 78. 
for gold, 79. 
indications of metals present. 78, 
losses in, 81. 
of matte, procedure for, 80. 
ores suited, 80. 
procedure, 75-77, 80, 124. 
Scorifiers, 19, 20. 
sizes, 74, 
spitting of, 79, 81. 
Segregation of metals in cooling, samp!- 
ing effeeted by, 120, 121. 
miliea 142. 
Silicates, classification of, 86. 
mixed, 89. 
Siliceous ores, assay of, 89-92. 
calculation of charge for, 89-92. 
Silver, effect in lead assay, 142. 
Slags, 85, 86. 
acid and basic distinguished, 88, 89. 
action of borax in, 86-88. 
for Class 1 basic ores, 92, 93. 
for Class 1 siliceous ores, 89-92. 


Slags, for Class 2 ores, 97, 98, 105, 106. 
fluidity of, 88. 
formation temperature, 88. 
Slag factors, bi-silicate, 91, 93. 
sub-silicate, 106. 
Slag forming constituents of ores, 84. 
Sodium carbonate, 2, 3. 
Solubility of metallic oxides in litharge, 
73, 74. 
Solutions, assay of, 133-139. 
Speiss, 8. 
Stack, height of, 12. 
size of, 12. 
support of, 12. 
Stanniferous: ores, crucible assay of, 
Ligeia 
Sulphuric acid combination method, 
for copper bullion, 123, 124, 126, 
127, 


Telluride ores, assay of, 112-114. 

Tellurium, effect of in cupellation, 113. 
in fusion, 113, 114. 

Thompson rider, 47. 

Tin ores, assay of, 116, 117. 


Vanning, 85. 
shovel, 85. 


Weighing, 41-43. 
by equal swings, 42. 
by method of swing, 42, 43. 
by ‘‘no deflection,” 43, 44. 
by substitution, 44. 
checking of, 44. 
gold, 69, 70. 
out ore, 76. 

Weights, 47, 48. 
calibration of, 48, 49. 


Zine effect in lead assay, 148. 














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